Extraction of lithium from primary and secondary sources ...

Hydrometallurgy 150 (2014) 192–208

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Review

Extraction of lithium from primary and secondary sources bypre-treatment, leaching and separation: A comprehensive review

Pratima Meshram a,b, B.D. Pandey a,⁎, T.R. Mankhand b

a CSIR—National Metallurgical Laboratory, Jamshedpur 831 007, Indiab Dept. of Metallurgical Engineering, IIT, BHU, Varanasi 221 005, India

⁎ Corresponding author.E-mail address: [email protected] (B.D. Pandey

http://dx.doi.org/10.1016/j.hydromet.2014.10.0120304-386X/© 2014 Elsevier B.V. All rights reserved.

a b s t r a c t
a r t i c l e i n f o
Article history:Received 27 October 2013Received in revised form 4 October 2014Accepted 12 October 2014Available online 23 October 2014

Keywords:LithiumPrimary oresBrinesLIBsHydrometallurgical recovery

In this comprehensive review resources of lithium and status of different processes/technologies in vogue orbeing developed for extraction of lithium and associated metals from both primary and secondary resourcesare summarized. Lithium extraction from primary resources such as ores/minerals (spodumene, petalite andlepidolite) by acid, alkaline and chlorination processes and from brines by adsorption, precipitation and ionexchange processes, is critically examined. Problems associated with the exploitation of other resources suchas bitterns and seawater are highlighted. As regards the secondary resources, the industrial processes followedand the newer developments aiming at the recovery of lithium from lithium ion batteries (LIBs) are describedin detail. In particular pre-treatment of the spent LIBs, leaching of metals from the cathode material in differentacids and separation of lithium and other metals from the leach liquors, are discussed. Although spent LIBs arecurrently processed to recover cobalt and other base metals but not lithium, there is a good prospect for therecovery of lithium in the coming years. Varying compositions of batteries for different applications requiredevelopment of a suitable recycling process to recover metals from all types of LIBs.

© 2014 Elsevier B.V. All rights reserved.

Contents

1. Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1932. Resources of lithium . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 193

2.1. Primary resources — minerals/clays and brines . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1932.2. Secondary resources — lithium ion batteries . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 194

3. Extraction of lithium from primary resources . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1953.1. Lithium extraction from minerals/clays . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 195

3.1.1. Acid process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1963.1.2. Alkaline process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1973.1.3. Chlorination process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1973.1.4. Other processes . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 197

3.2. Lithium extraction from brines/sea water/bitterns . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1973.2.1. Adsorption process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1983.2.2. Precipitation process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1993.2.3. Ion exchange/Solvent extraction process . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 199

3.3. Extraction of lithium from secondary resources — lithium ion batteries . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2003.3.1. Major industrial processes . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2003.3.2. Recent development in recycling of lithium ion batteries . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 202

4. Conclusions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 206Acknowledgements . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 206References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 206

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Table 1Principal commercial lithium minerals with compositiona.

Mineral Formula % Lithium content

Theoretical Range in commercialminerals

Spodumene LiAlSi2O6 or Li2O·Al2O3·4SiO2 3.73 1.9–3.3Lepidolite LiKAl2F2Si3O9 or

LiF·KF·Al2O3·3SiO2

3.56 1.4–1.9

Amblygonite LiAlFPO4 or 2LiF·Al2O3·P2O5 4.74 3.5–4.2Triphylite LiFePO4 or Li2O·2FeO·P2O5 4.40 2.5–3.8Petalite LiAlSi4O10 or Li2O·Al2O3·8SiO2 2.27 1.6–2.21Bikitaite LiAlSi2O6·H2O 3.28 1.35–1.7Eucryptite LiAlSiO4 5.53 2.34–3.3Montebrasite Li2O·Al2O3·2SiO2 3.93 0.9–1.8Jadarite LiNaSiB3O7(OH) 3.39 0.096–0.1Zinnwaldite LiKFeAl2F2Si3O10 or

LiF·KF·FeO·Al2O3·3SiO2

1.7 1.21–1.3

Hectorite Na0.3(Mg,Li)3Si4O10(F,OH)2 0.56 0.36Zabuyelite Li2CO3 18.75 –

a IndustrialMinerals andRocks (2006), Norton and Schlegel (1955), Schaller (1937), andSiame and Pascoe (2011).

193P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

1. Introduction

Lithium is the 25th most abundant element (at 20 mg/kg) in theearth's crust. Lithium finds an application in rechargeable lithium ion bat-teries (LIBs) because of its very high energy density by weight and highelectrochemical potential (3.045 V). With a present consumption levelof ~22% of the total lithium produced in LIBs, it is expected to reach to~40% by 2020 (Wang et al., 2012). Besides batteries, at present it hasmajor applications in glass and ceramics (30%), greases (11%),metallurgi-cal (4%) industries and also in chemicals/pharmaceuticals, rubbers etc.(Garrett, 2004; Holdren, 1971). As per the Madrid Report of July 2010,lithium falls in the border-line of low-to-medium in supply and demand,primarily due to scanty resources (Tedjar, 2013). It is also understood thatthe demand for lithium is increasing further due to its application in nu-clear and strategic areas. As per recent estimates of lithium reserves, outof 103 deposits with more than 1,00,000 t of lithium, each deposit has adifferent mineralogical composition and therefore, requires the appropri-ate technology to process (Gruber et al., 2011). In view of this, lithium isusually extracted from its mineral that is found in igneous rocks (chieflyspodumene) and from lithium chloride salts found in brine pools, whileignoring other resources including the low-grade ores.

The rising demand for lithium for various applications thus calls forprospecting and processing all viable resources. Lithium extractionfrom ores/minerals utilizes roasting followed by leaching, while itsextraction from brines includes evaporation, precipitation, adsorptionand ion exchange (Garrett, 2004). Lithium can be extracted from LIBsby leaching followed by precipitation, ion exchange or solvent extrac-tion and electrolysis (Shuva and Kurny, 2013). It is estimated that 250t of ore (spodumene) or 750 t of brine or 28 t of lithium ion batteriesof mobile phones and laptops or 256 batteries of electric vehicles(EVs) are required to produce 1 t of lithium (Tedjar, 2013). Lithium con-centrate obtainedmainly by the flotation of pegmatites (ore), is pulver-ized and leached in hot acid, and lithium is precipitated as lithiumcarbonate (Tahil, 2010). The processing of pegmatites is expensive ascompared to that of the brines due to the heating and dissolutionsteps involved, but the higher metal concentration in pegmatites partlycompensates for the cost. Because of the cost factor in the lithiumextraction from brine compared to the ores, many deposits of spodu-mene are not currently being mined/processed. Lithium is also presentin seawater, but the concentration is too low to be economical. Asregards lithium metal, it can be produced by both carbothermic reduc-tion and metallothermic reduction of oxide (at times hydroxide) andalso by electrolysis of LiCl (Kipouros and Sadoway, 1998). In viewof the scattered literature on the extraction of lithium from primaryresources viz., ores, minerals and the brines, it is consideredworthwhileto review the details and discuss critically the merits and demerits ofvarious processes in vogue or being developed.

With the increasing use of LIBs in mobile phones, laptops, cam-corders, tracking systems, military and medical devices and in largeenergy storage systems including that of transportation applications(Nonemillion EVs expected by 2015), therewill be a significant pressureon lithium resources and its supplies (Kuo, 2011). The disposal of spentbatteries may involve landfilling, stabilization, incineration or recycling.In landfills, heavymetals have the potential to leach slowly into the soil,groundwater or surface water.

The methods for recycling spent LIBs are based mainly on pyro-/hydro-metallurgical processes (Li et al., 2009a). The disadvantage of allpyro-recycling processes is that lithium is not recovered. The traditionalpyrometallurgical processes can burn off all the organic electrolyte andbinder, and facilitate the leaching of valuablemetals. In the hydrometal-lurgical processes, the dismantled electrodes are dissolved in concen-trated acid and the metal rich leach solutions are treated to recoverthe individual metal by the different methods mentioned above. Theseprocessesmay producewastewater containing fluoridewhich is difficultto treat, and can pollute the environment due to the incompleterecycling of the organic binder and electrolyte. There is an inconsistent

policy about the fate of discarded lithium ion batteries in e-waste thatis distributed internationally. Lithium batteries also contain potentiallytoxic materials including metals, such as copper, nickel and lead, andorganic chemicals, such as toxic and flammable electrolytes containingLiClO4, LiBF4, and LiPF6. Defunct Li-ion batteries are classified as hazard-ous due to their lead (Pb) (6.29 mg/kg), cobalt (163,544 mg/kg), copper(98,694 mg/kg) and nickel (9525 mg/kg) contents with exceeded limitsof chromium, lead, arsenic and thallium (Bernardes et al., 2004; Kanget al., 2013). Human and environmental exposures to these chemicalsare typically regulated during the manufacture of lithium batteriesthrough occupational health and safety laws, and potential fire hazardsassociated with their transportation. These findings support the needfor stronger government policies at the local, national, andinternational levels to encourage recovery, recycling, and reuse of lithi-um battery materials. In view of the above, efforts must be made todevelop an environmentally benign and economically viable technologyfor recycling spent LIBs.

This review focuses on the primary and secondary resources oflithium available for exploitation and provides comprehensive detailson the conventional/currently practiced lithium extraction methodsvis-a-vis the resource type. The resources that are covered includeores/minerals/clays and brines/seawater and bitterns, and lithium ionbatteries for the hydrometallurgical recovery of lithium.

2. Resources of lithium

2.1. Primary resources — minerals/clays and brines

Lithium is produced from a variety of natural sources, e.g., mineralssuch as spodumene, clays such as hectorite, salt lakes, undergroundbrine reservoirs etc. Lithium is a minor component of igneous rocks,primarily granite. Themost abundant lithiumcontaining rocks/mineralsare pegmatites, spodumene and petalite. Other minerals are lepidolite,amblygonite, zinnwaldite and eucryptite (Ferrell, 1985). Zinnwalditeis the impure form of lepidolite with higher content of FeO (up to11.5% Fe as FeO) andMnO (3.2%) (Paukov et al., 2010). Pegmatites con-tain recoverable amounts of lithium, tin, tantalum, niobium, berylliumand other elements. Table 1 lists the principal commercial lithiummin-erals found in pegmatites along with their composition. The theoreticallithium content in these minerals is 3% to 5.53%, but most mineral de-posits have around 0.5%–2% Li and the pegmatite-bearing ores that areoften exploited have b1% Li (Mohr et al., 2010). Spodumene is the pri-mary lithium mineral being mined.

Among the clay minerals, hectorite, a type of smectite is rich in lith-ium and magnesium, and generally contains 0.3 to 0.6% Li. The best

Table 2Geographical availability for various lithium resources and their lithium contents.

Resources Li content % (w/w) Location of the largest amount References

PegmatitesSpodumene 1.9–3.3 Australia Industrial Minerals and Rocks (2006), Legers (2008),

Clarke (2013), AMRA (2013)Lepidolite 1.53–3.6 ZimbabweZabuyelite 17–18.75 ChinaPetalite 1.6–2.1 ZimbabweAmblygonite 3.5–4.2% Canada, Brazil, SurinamEucryptite 2.34 Zimbabwe

Sedimentary rocksSmectite (hectorite) 0.27–0.7 Hector, California. and Nevada, US Industrial Minerals and Rocks (2006)Lacustrine evaporites 1.8% lithium oxide Jadar Valley, Serbia

BrineContinental b2700 mg/L (with K, B) Chile Evans (2008,2009),

Legers (2008),Clarke (2013),Mohr et al. (2012)

~532 mg/L BoliviaGeothermal b400 mg/L (with Mn and Zn) United StatesOil field b700 mg/L (with Br) United StatesSea water 0.18 mg/L Chile, Argentina, China and Tibet

Secondary resourceSpent LIBs 2–7% – Shuva and Kurny (2013)

194 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

known hectorite deposit with 0.7% Li is in Hector, California. Flint claysand other high-alumina clays contain b0.01 to 0.5% Li. Jadarite is anewly discovered lithium-boron containing mineral found in Serbia(Mohr et al., 2012). These minerals are often concentrated to around2%–4% Li for use in the ceramics and glass industry (Garrett, 2004).

Seawater contains about 0.1–0.2 mg/L Li (Bach and Wasson, 1981).Total amount of metallic lithium in seawater (globally) is estimated tobe ~230 Gt. Brine sources include lithium found in salt water deposits—lakes, salars, oilfield brines, and geothermal brines. Oilfield brines areunderground brine reservoirs that are locatedwith oil. Geothermal brinesare underground brines naturally heated, e.g., in the Salton Sea California.Brines containing lithium make up 66% of the world's lithium resources;pegmatites make up 26% and sedimentary rocks make up 8% (Gruber,2010; Kesler et al., 2012).

Almost 70% of the global lithium deposits are concentrated in SouthAmerica's ABC (Argentina, Bolivia and Chile) region. Table 2 details thegeographical availability of various resources of lithium vis-a-vis lithiumcontent and their locations. The lithium concentrations in the salars ofChile, Argentina, and Bolivia are in the range 0.04–0.16%. According toYaksic and Tilton (2009) the resource of lithium is estimated to be64 Mt. Chile has the world's largest resource of brine (7.5 Mt, 1500–2700 mg/L Li) containing lithium, followed by Bolivia (resource: 9.0 Mtwith 532 mg/L Li) and Argentina (resource: 2.6 Mt, 400–700 mg/L Li)and these three countries account for almost 80% of the world's brinereserves (Mohr et al., 2012). Estimates of lithium resources are publishedextensively (Clarke and Harben, 2009; Gruber et al., 2011; Ono, 2009;Evans, 2010a, 2010b; USGS, 1980, 1986, 2005, 2009, 2010, 2011, 2012,2013).

Lithium rock production began with lithium minerals (1899) in theUSA (Garrett, 2004). Since the first lithium production from brines atSearles Lake, USA in 1936, brines are exploited largely in South Americaand China. The largest producer of lithium in the world is Chile wherelithium is extracted from salar brine at the Atacama Salt Flat. Lithiumproduction from brines is also at salt lakes in Tibet and Qinghai inChina, besides at Nevada in the United States. Several newer installations(by 2013) are on various stages of exploration/operation for brine sourcewhich are: one more in China, six in Argentina, three in Chile and one inBolivia (Clarke, 2013). Currently 8% of lithium is obtained from salt lakebrines and sea by sedimentation. Significant quantities of lithium com-pounds and ore concentrates are also produced in Australia, Canada,Portugal, Russia and Zimbabwe. Currently, Australia produces lithiumconcentrate from spodumene at the mines in Mt Catlin, WesternAustralia. The 160,000 t/annum concentrate produced in WesternAustralia is processed to 16,000 t/annum lithium carbonate in its Chinese

plant. Lithium carbonate is mostly produced from both ores and brinesand the production figures are often expressed as lithium carbonateequivalent (LCE). Other chemicals such as lithium chloride and hydrox-ide are also produced in varying amounts.

In India small pocket deposits mostly comprising of lepidolite inpegmatites of mica fields are located. The maximum lithium content(lepidolite with 2–6% Li2O) is found in Jharkhand followed by that ofChhattisgarh (2.56% Li2O) and Rajasthan (2.25% Li2O as pegmatites inUdaipur, Bhilwara, Jodhpur and Ajmer, and zinnwaldite in Dagana).Others include spodumene in Raichur, Karnataka (Banerjee et al., 1994),amblygonite in granitic rocks of Paddar (Kashmir) and lepidolite atDhir-Bil (Goalpara), Assam. Lithium bearing bauxite has also been identi-fied in the Salal area, Jammu (Brown and Dey, 1955; Krishnaswamy,1979; Roonwal et al., 2005).

2.2. Secondary resources — lithium ion batteries

Out of the various secondary resources, spent LIBs are themost prom-inent secondary source of lithium and other metals. To recycle thesematerials, it is desired to understand the construction/composition ofthe cells in brief and how they transform during their use.

Lithium ion battery is a term generally used for a battery which haslithium metal, lithium alloy or material adsorbing lithium ions for itsnegative active material. LIB uses carbon as an anode and lithium ionsexist in the carbon material; there is no metallic lithium at any state ofcharge during normal usage. Depending on their technical constructionand properties batteries are categorized as either primary or secondary.From the legislative viewpoint batteries are also categorized as portable(household) and vehicle or industrial batteries. Primary cells are con-structed with metallic lithium. The metallic lithium in a non-rechargeable primary lithium battery is a combustible alkali metal thatself-ignites at 178 °C, and when exposed to water/seawater reacts exo-thermically and releases hydrogen. Primary batteries are single-use asirreversible discharge reactions occur in the cells and after use theyare disposed off. Secondary battery cells have a chemistry that allows re-versing the discharge reaction and are rechargeable. LIBs are of the re-chargeable secondary type. The functional parts of LIBs are the cathode,anode, electrolyte and separator, which are housed in a protectivemetal casing.

The chemical reaction in the cell expressed below shows the formsin which lithium and cobalt can be present in the spent LIBs.

LiCoyOz þ 6C→LixC6 þ Li1−xCoyOz: ð1Þ

Table 3Various components of LIBs, materials used and their meritsa.

Component Wt.% of the totalweight of the battery

Material Structure Properties/Merits

Cathode 39.1 ± 1.1 LiCoO2 Layered High structural stability and can be cycled for N500 times with 80–90%capacity retention

LiMn2O4 Spinel Attractive for ecological and economic reasons; discharges ~3 VLiNiO2 Layered Cheaper & possesses higher energy density (15% higher by volume, 20% higher

by weight), but less stable & less ordered as compared to LiCoO2

LiFePO4 Olivine Suitable for biomedical applications because of higher safety levels andlower costLi2FePO4F Olivine

LiCo1/3Ni1/3Mn1/3O Layered/spinel Possesses high capacity with structural and thermal stability, and safe to useLi(LiaNixMnyCoz)O2 Layered/spinel

Anode Carbon Graphite Low cost and availability. It has the ability to reversibly absorb and release largequantity of Li (Li:C = 1:6)Hard carbon Microspheres

Electrolyte Lithium salt like LiPF6, Li[PF3(C2F5)3]or LiBC4O8 in organic solvents

Withstands high temperature and possesses high mobility of Li ions

Plastic case 22.9 ± 0.7 Polyethylene terephthalate layers, a polymer layer and apolypropylene layer, layers of carbonized plastic

Hermetically sealed battery body which converts chemical energy to electricalenergy in order to generate current

Outer casing 10.5 ± 1.1 Stainless steel, aluminiumCopper foil 8.9 ± 0.3 Copper ~14 μm thickAluminium foil 6.1 ± 0.6 Aluminium ~20 μm thickPolymer foil &electrolyte

5.2 ± 0.4 Polyethylene, polypropylene or compositepolyethylene/polypropylene films

Use of 3–8 μm layers (PP/PE/PP) with 50% porosity

Solvent 4.7 ± 0.2 Ethylene carbonate, dimethyl carbonate& diethyl carbonate

Non-aqueous

Electrical contact 2.0 ± 0.5 Aluminium and copper Conductive

a Wakihara and Yamamoto (1998), Gaines and Cuenca (2000), Nazri and Pistoia (2004) Paulino et al., (2008) and Fergus (2010).


The activemass (cathode, anode and electrolyte) of LIBs comprises ofalmost 40% of the weight whereas ~30% (wt) of these components arecarbon (anode). A number of chemical species in the cathode materialwithin the lithium-ion family have been reported including the type,structure of the materials used and their electrochemical propertieswhich are listed in Table 3. Generally LIBs have a short life of 2–3 yearswhether they are used or not. The spent batteries are a good source ofseveral metals like Li, Co, Ni, Co, Mn etc.

3. Extraction of lithium from primary resources

3.1. Lithium extraction from minerals/clays

Lithium is extracted from its minerals by two processes — acid andalkaline though chlorination is also attempted in some cases. Averill

Table 4Lithium extraction from its minerals/clays by acid process.

Raw material % Li Experimental conditions

Pre-treatment Roastingtemp. (°C)

Time(h)

Water leac

L/S ratio

Lepidolite conc. 2.0 Sulfation roasting 850 0.5 2.5:1

Lepidolite conc. 2.0 Salt roasting withNa2SO4 + CaCl2

880 0.5 0.8:1

Lepidolite conc. 2.55 Sulfation roastingNa2SO4:Li molarratio = 2:1

1000 0.5 15:1

Lepidolite conc. 1.79 Roasting with ironsulfate

850 1.5 1:1

Zinnwaldite conc. 0.96 Roasting withsodium sulfate

850 1 10:1

Petalite/Lepidolite 1.9 Calcination 1100 2 7.5/1Roasting (H2SO4) 300 1

Spodumene 4.21 Calcination 1100 1 Leaching &precipitatiRoasting (H2SO4) 250 0.167

Spodumene 2.81 Roasting & H2SO4

leaching1050–1090 0.5 1:4

Montmorillonite clay 1.2% Li2O Without roasting – – 5:1 (H2SO

and Olson (1978) have reviewed methods and techniques for theextraction of lithium from ores, brines and clays. Processes followedfor the extraction of lithium from different resources have also beencompiled in detail by Garrett (2004).

In order to process ores/concentrates, acid digestionwith H2SO4maybe followed for decomposition of the silicate structure at 250–400 °Cwhich is suitable for the processing of lepidolite, amblygonite andzinnwaldite (Kondás and Jandová, 2006). In the sulfate process lithiumminerals such as lepidolite are decomposed at high temperature in thepresence of potassium and/or sodium sulfate. Alkali digestion is suitablefor the decomposition of spodumene and lepidolite largely by the treat-ment of potassium carbonate to produce lithium hydroxide. In the alka-line/gypsum process the mineral is reacted with limestone or a mixtureof calcium sulfate with calcium oxide and/or hydroxide by heating toconvert silicate to soluble lithium aluminate from which LiOH or Li2CO3

% LiExtraction

Li2CO3

purity(%)

References

hing

Temp. (°C) Time (h)

Room temp.(RT)

0.5 91.6 – Yan et al.(2012a)

RT 0.5 ~90 N99.5 Yan et al.(2012b)

85 3 90.4 – Luong et al.(2013)

RT 1 85 (opensystem)93 (closedsystem)

– Luong et al.(2014)

RT 0.5 N90 N90 Siame andPascoe (2011)

50 1 97.3 99 Sitando andCrouse (2012)

on– – 90 – Mcketta (1988),

Tahil (2010)225 1 96 ~99.6 Chen et al.

(2011a, 2011b),Clarke (2013)

4 leach) 250 1.5 90 Li2SO4 Amer (2008)

Table 5Lithium extraction from its minerals/clays by alkali process.

Raw material % Li Pre-treatment Experimental conditions % Liextraction

Li2CO3

purity (%)References

Roasting temp.(°C), time (h)

Water leaching

Time (h) Temp. (°C) (L/S ratio)

Lepidoliteconcentrate

1.4 Defluorination 860, 0.5 h 1 150 4:1; lime: defluor.lepidolite: 1

98.9 99.9 Yan et al. (2012c)

Zinnwalditeconcentrate

1.4 Gypsum roast 950, 1 h 0.167 90 10:1 96 99 Jandová et al.(2009)

Zinnwaldite(tailing sample)

2.07%Li2O

Gypsum roast 1050 0.5 85 10:1 84 –


0.19 Roasting withgypsum & Ca(OH)2

975 1 90 5:1 93 – Kondás and Jandová(2006)


1.21 Roasting with CaCO3 825, 1 h 1 90–95 5:1 ~90 99.5 Jandová et al. (2010)


1.29 Roasting with CaCO3 825 4 95 10:1 84 – Vu et al. (2013)

Montmorilloniteclay–hectorite

0.3–0.6 Roasting withCaCO3

CaCO3–CaSO4

750900

– RT – 80 99 Lien (1985), Crockerand Lien (1987)


is obtained. Thus most of the alkaline processes involve either theheating of lithiumminerals with alkali salts or in more advanced hydro-thermal processes by decomposition in solutions containing Na2CO3,NaOH, Na2SO4 and/or other alkali salts at elevated temperature andpressure. Ion-exchange processes are applied for the processing of theleach liquors obtained from spodumene, petalite and partly fromzinnwaldite. Tables 4–6 summarize recent work on lithium extractionfrom its minerals and clays by using different approaches.

3.1.1. Acid processSulfation roasting of lepidolite followed by water leaching was

recently reported by Yan et al. (2012a). The lithium extraction effi-ciency of 91.6% could be achieved at a mass ratio of lepidolite/Na2SO4/K2SO4/CaO of 1:0.5:0.1:0.1 and roasting at 850 °C (Table 4). Roasting at880 °C with a mass ratio of lepidolite/Na2SO4/CaCl2 of 1:0.5:0.3 resultedin improved recovery (~95% Li) of lithium (Yan et al., 2012b). Above99.5% pure lithium carbonatewas obtained by evaporation and precipita-tionwithNa2CO3. Luong et al. (2013) examined the sulfation roasting of aKorean lepidolite ore with sodium sulfate at 1000 °C, while achievingextraction of ~90.4% Li. Recently, the roasting of lepidolitewith iron sul-fate at 850 °C and water leaching were investigated by Luong et al.(2014). Leaching of the calcines obtained from the open and closed sys-tems yielded leach liquors containing ~7.9 g/L Li and ~8.7 g/L Li, corre-sponding to the extraction of ~85% and ~93% Li, respectively.

Lithium extraction (N90%) by roasting a low grade zinnwalditeconcentrate (0.96% Li) with sodium sulfate followed by water leaching,was described by Siame and Pascoe (2011). A study by Sitando andCrouse (2012) showed the extraction of 97.3% Li (~5 g/L) by roastinga Zimbabwean petalite ore concentrate (4.1% Li2O) with concentratedsulfuric acid at 300 °C followed by water leaching at 50 °C. The leachliquor was evaporated till lithium was concentrated to N11 g/L and

Table 6Lithium extraction from its minerals/clays by chlorination process.

Raw material Experimental conditions

Roasting temp. (°C) Time (h) Leaching

L/S ratio

Lepidolite 935 13 –

Spodumene (3.58% Li) 1000 4 10:1Lepidolite concentrate (2.0% Li) 880 0.5 2.5:1Spodumene (7.2% Li2O) 1100 2.5 –

Lepidolite ore (3.70% Li2O) With CaCO3 + CaCl2, 950 5 10:1

Li2CO3 of ~99% purity was precipitated by the addition of Na2CO3 at95–100 °C. Amer (2008) reported the extraction of 90% Li from theEgyptian montmorillonite type clay containing lithium in 90 min ofreaction in sulfuric acid at 250 °C.

In order to process spodumene it is desired to convertα-spodumeneto β-phase by roasting at 1070–1100 °C (Chen et al., 2011b; Clarke,2013; Tahil, 2010). Tahil (2010) reported the roasting of spodumenein a kiln at ~1100 °C. The calcine was mixed with sulfuric acid androasted at 250 °C and then leached inwater to yield a solution of lithiumsulfate. Reaction of β‐spodumene with H2SO4 is shown as reaction (2)(Mcketta, 1988):

Li2O⋅Al2O3⋅4SiO2ðsÞ þ H2SO4ðconcÞ→Li2SO4ðsÞ þ Al2O3⋅4SiO2ðsÞþ H2OðgÞ: ð2Þ

Lithium carbonate can be recovered by the addition of sodium carbon-ate to the solution after pH adjustment, purification and evaporation(Reaction (3)).

Li2SO4ðaqÞ þ Na2CO3ðaqÞ→Li2CO3ðsÞ þ Na2SO4ðaqÞ: ð3Þ

The world's first continuous plant to convert spodumene concentrateto lithium carbonate by calcination, roasting of calcine with H2SO4 andwater leaching, was commissioned in 2012 by Galaxy Resources inChina (Clarke, 2013).

One of the drawbacks of the sulfuric acid method to treat lepidolite,petalite and zinnwaldite is the requirement of a high concentration ofacid and complex purification processes, whereas spodumene needsto be converted to the more leachable β‐phase at higher temperature.

% Li extraction Product References

Temp. (°C) Time (h)

80 1 ~100 LiCl Löf and Lewis (1942)80 1 58 LiCl Peterson and Gloss (1959)60 0.5 92.8 Li2CO3 Yan et al. (2012d)– – 100 LiCl Barbosa et al. (2014)90 1 80.0 LiCl Vyas et al. (1975)


3.1.2. Alkaline processIn the alkaline process (Table 5), spodumene or lepidolite ore concen-

trates are ground and calcined with limestone at 825–1050 °C. Theresulting calcine is crushed, milled and treated with water to yield lithi-um hydroxide which can be converted to chloride by reaction with hy-drochloric acid. The recovery in this method is approximately 85–90%(Averill and Olson, 1978). The reaction during calcination of spodumenewith limestone can be represented as :

Li2O⋅Al2O3⋅4SiO2ðsÞ þ CaCO3→Li2OðsÞ þ CaO⋅Al2O3⋅4SiO2ðsÞþ CO2ðgÞ: ð4ÞLepidolite was pre-roasted at 860 °C under water steam atmosphere

for defluorination followed by pressure leaching of the defluorinatedmass in a lime–milk autoclave at 150 °C. In this process 98.9% lithiumwas extracted (Yan et al., 2012c). During the roasting with steam forma-tion of phases such as lithium aluminium silicate (beta-eucryptite) andleucite (KAlSi2O6) was noticed (Karstetter, 1971) as per reaction (5)

2LiF⋅KF⋅Al2O3⋅3SiO2 þ 4H2O→Li2O⋅Al2O3⋅2SiO2 þ 2KAlSi2O6þ 4H2 þ 2F2: ð5ÞExtraction of lithium from a zinnwaldite containingwastewas inves-

tigated by Jandová et al. (2009). The concentrate (1.40% Li) obtainedfrom dry magnetic separation of the waste (0.21% Li) was treated bythe gypsum roast method (with CaSO4 and Ca(OH)2). About 96% Li wasleached out from the sinter made at 950 °C. Earlier the concentrateprepared from this waste by gravity and dry magnetic separation, wasroasted by the gypsum method at 975 °C and water leached to recover~93% Li (Kondás and Jandová, 2006). In another study Jandová et al.(2010) reported that the roasting of the above concentrate withCaCO3·Li2CO3 with 99.5% purity was separated from the leach liquorsby either converting the alkaline liquor to the carbonated solution byCO2 bubbling or by solvent extraction with LIX54 and TOPO as anextractant followed by stripping with diluted H2SO4; the first methodprovided a higher yield. Siame and Pascoe (2011) obtained recovery of~84% Li from the roasted zinnwaldite concentrate at 1050 °C with lime-stone, gypsum and sodium sulfate. A similar process was reported by Vuet al. (2013) from zinnwaldite by sinteringwith CaCO3 powder at 825 °C,followed by water leaching and precipitation.

Crocker and Lien (1987) reported the recovery of N85% Li by roastingmontmorillonite type clays (0.3–0.6% Li) with KCl–CaSO4 or CaCO3–

CaSO4 followed by water leaching. Lithium silicate in the clay was con-verted to Li2SO4 by roasting a pelletized mixture of clay, limestone andgypsum at 900 °C in a direct gas-fired rotary roaster (Lien, 1985).Formation of lithium sulfate may be represented by reactions (6) and (7)

CaSO4 � 2H2Oþ SiO2→CaSiO3 þ SO2 þ12O2 ð6Þ

Li2Si2O5 þ SO2 þ12O2→Li2SO4 þ 2SiO2: ð7Þ

The initial product obtained from the alkaline process is lithiumhydroxide which can be converted to carbonate or chloride salt.

3.1.3. Chlorination processA less commonmethod is the chlorination process in which the ore is

roasted in the temperature range880–1100 °C in the presence of chlorinegas or HCl depending upon the type of mineral treated (Table 6). Forinstance chlorination of lepidolitewithHCl can be achieved at lower tem-perature (935 °C) while giving a high yield (~100%) of lithium duringleaching as compared to the roasting of spodumene (Löf and Lewis,1942). Lithium in spodumene can be converted to LiCl almost quantita-tively at higher temperature (1100 °C) in 2.5 h with Cl2 gas (Barbosaet al., 2014); the recovery is found to be quite low (58%) at the lowertemperature (1000 °C) (Peterson and Gloss, 1959). In the chlorinationprocess, when the ore is sintered with NH4Cl and CaCl2 in a furnace at

750 °C, ~98% of the lithium contained in spodumene is converted to itschloride which can be water leached (Zelikman et al., 1996). Vyaset al. (1975) reported a similar process using CaCO3 and CaCl2 to roastIndian lepidolite at 950 °C by which 80% Li was recovered as LiCl. Thechlorination process with calcium chloride and/or sodium chloridewith lepidolite (M = Li, K, Rb, Cs) may be represented as:

2NaCl þ 6SiO2 þM2O þ Al2O3→2NaAlSi3O8 þ 2MCl ð8Þ

CaCl2 þ SiO2 þ M2O→CaSiO3 þ 2MCl ð9Þ

CaCl2 þ 2SiO2 þ Al2O3 þM2O→CaAlSi2O8 þ 2MCl: ð10Þ

Chlorination roasting with a mixture of CaCl2 and NaCl gives a betterlithium extraction yield because of its lower melting point than eitherof the agents, which increases the fluidity of chloride melt. This allowsdiffusion of the chlorinating agent to the surface of the lepidolite facili-tating the lithium extraction selectively and yielding a pure product.Crocker and Lien (1987) also reported a process for selective chlorina-tion of hectorite (0.3–0.65% Li) in clays with limestone at 750 °C using20 wt.% HCl.

3.1.4. Other processesChlorination roasting of ores requires corrosion-resistant equip-

ment. To overcome this drawback the autoclave method is used. Chenet al. (2011b) reported a process to treat β-spodumene obtained inthe calcination ofα-spodumene, by sodiumcarbonate solution in an au-toclave at a liquid/solid ratio of 4 and Na/Li of 1.25 at 225 °C. Duringpressure leaching lithium carbonate and analcime slurry are formedaccording to reaction (11).

β−Li2O⋅Al2O3⋅4SiO2 þ Na2CO3 þ nH2O→Li2CO3þ Na2O⋅Al2O3⋅4SiO2⋅nH2O: ð11Þ

The slurry was leached in carbon dioxide to form lithium bicarbonatewhich on heating produced lithium carbonate of 99.6% purity as perreactions (12) and (13).

Li2CO3 þ CO2 þ H2O→2LiHCO3 ð12Þ

2LiHCO3→Li2CO3 þ CO2 þ H2O: ð13Þ

Medina and El-Naggar (1984) developed an alternate method ofchlorination to treat spodumene with a recovery of ~87% Li by roastingat 1150 °Cwith amixture of 8:1 (wt.) tachyhydrite:spodumene follow-ed by water leaching. Because the production of lithium carbonate fromspodumene is energy intensive and expensive, lithium carbonate ismostly produced from the brines.

3.2. Lithium extraction from brines/sea water/bitterns

In order to meet the growing demand of lithium, brines and bitternshave received increasing attention. Tables 7 and 8 elaborate the summa-ry of work carried out recently for the extraction of lithium from seawater/brines/bitterns. Production process for brine-water lithium costs30% to 50% less than that of the mined ores (Abe, 2010). As mentionedearlier, lithiumcarbonate is produced frombrines by an evaporative con-centration and refining method. Firstly, brine is concentrated by solarevaporation over a year in a pond to crystallize sodium, potassium andmagnesium chlorides. During the refining process calcium carbonateis roasted and then added to the solution of LiCl for the removal of

Table 7Adsorption process for lithium extraction with product synthesized from seawater, brines and bitterns.

Source/Raw material Adsorbent used Conditions Adsorption/Remarks Product References

Seawater H1.6Mn1.6O4 Adsorbent: 200 mg, sea water: 50 L,4 weeks; desorption by 0.5 M HCl, 1 day

Max. uptake: 40 mg Li/g adsorbent – Chitrakar et al.(2001)

Seawater (Li 0.17 mg/L) Spinel type Mn-oxide 30 °C, 15 days b90% Li,Li uptake: 10.6 mg/g adsorbent

LiCl Umeno et al.(2002)

Seawater (Li 0.15 mg/L) λ-MnO2 (granulated) 150 days Recovery—264 g LiCl in 791 g driedprecipitated salt (816 m3 seawater)

LiCl Yoshizukaet al. (2006)

Artificial seawater(Li 0.2 mg/L, 8.01 pH)

Ion-sieve type Mn-oxidespinels: HMg0.5Mn1.5O4(I)HZn0.5Mn1.5O4 (II)

Adsorption at 0.4 M HCl, 5 days Adsorption: 88% Li by adsorbent I and 89% Liby adsorbent II.Equil. sorption: 30.3 mg/g (I) and33.1 mg/g (II).

LiCl Chung et al.(2004, 2008)

Seawater (1 mmol/L Li) HMn2O4 60 °C, 24 h Loading capacity: 1.53 mmol/g sorbent Li salt Wajima et al.(2012)

Brine (Salars de Uyuni,Bolivia)

LimMgxMnIIIyMnIV

zO4

(0 b x ≤ 0.5)24 h, pH 6.5 Adsorption capacity: 23–25 mg/g

adsorbentLi2CO3 Chitrakar et al.

(2013)Seawater (Li0.192 mg/L)

Surface deposition oncorrosion product of Al

30 °C, 10 days 34% Li – Takeuchi(1980)

Salt lake bitterns Hydrated alumina:LiOH = 2 M ratio

pH 5.8 Adsorption:0.6–0.9 mg/g sorbent – Dong et al.(2007)

Egyptian bitterns(19.5, 5.5, 8.8 mg/L)

Al(OH)3 30 °C, pH 9 Adsorption capacity: 123 mg/g adsorbent LiAlO2 Hawash et al.(2010)

Brine (Salar de HombreMuerto, Argentina)

Hydrated alumina LiCl solution—1% Li (20 times conc.) by solar evaporation.

Adsorbed Li eluted by acid andprecipitated by sodium carbonate

Li2CO3 Clarke (2013)


Mg(OH)2. A general flowsheet of Li2CO3 production from brine water isshown in Fig. 1.

The basic approaches for the separation of mineral products (K, Mg,Na, Ca, Li) from the seawater comprises of flotation (using anionic col-lectors), sorption, ion exchange, solvent extraction etc. as described byKoyanaka and Yasuda (1977). The existing evaporation process for lith-ium recovery from brine lakes is time consuming and suffers from lowrecovery efficiency. Besides, tremendous burden is posed on the envi-ronment due to waste generation and substantial water consumption.

3.2.1. Adsorption processVarious types of adsorbents have been used for selective lithium

recovery from seawater and brines. In the adsorption method certaininorganic ion-exchangers such as the spinel-type manganese oxideshow extremely high selectivity for lithium from seawater (Kitajouet al., 2003; Ooi et al., 1989, 1991; Umeno et al., 2002; Yoshizuka et al.,2002, 2006). Suchmaterials exhibit high adsorption capacities in alkaline

Table 8Precipitation and other processes for lithium extraction from seawater and brines.

Source/Raw material Process Conditions

Synthetic solution,Geothermal water(Li = 10 mg/L)

Precipitation pH 12.5 for [A

Synthetic solution (2.5 MLiCl, 0.3 M CaCl2 and 0.15 MMgCl2)

Precipitation followed by IX using Poly BD®R45HTLO (MC50), Lewatit® (TP 207),Dowex®(Y80)

Precipitation wat 80 °C.IX—50 °C, 30

Seawater (0.12–0.16 mg/L Li) Integrated ion-exchange method 1st stage: sorp150 days(264 g LiCl in2nd stage: SepSr(II) and Mn((pH 9)

Seawater (0.18–0.20 mg/L Li) Two stage precipitation 1st stage pH :2nd stage: Na

Uyuni Salar brine, Bolivia,15–18 g/L Mg,0.7–0.9 g/L Li

Two stage precipitation with lime& Na-oxalate

1st stage pH:2nd stage: sod80–90 °C

Brine from Salar de HombreMuerto, Argentina

Precipitation (2.5 g/L LiCl from solar pond)with lime and Na2SO4

Separation ofCa as sulfate

Brine (high Mg/Li ratio) Electrochemical Electrolyte: 0

medium (pH of seawater being ~8) for Li+ in the presence of alkali andalkaline earth ions. For instanceKitajou et al. (2003) reported the separa-tion of Li+ from a large amount of Na+ by the spinel-type λ-MnO2

whereby Li+ was concentrated 400 times leaving most Na+ in the sea-water. Extraction/separation of lithium from brines and such resourcesis summarized in Table 7.

The manganese oxide (H1.6Mn1.6O4) prepared from the precursor,Li1.6Mn1.6O4 by hydrothermal and reflux methods, showed the maxi-mum uptake of 40 mg Li/g of adsorbent from the seawater, the highestamong the inorganic adsorbents (Chitrakar et al., 2001). The very finesize (nano-size range) of the synthesized manganese oxide was foundto be responsible for its high adsorption capacity towards lithium ascompared to other adsorbents. Adsorption of lithium from seawater bya spinel type λ-MnO2 produced low purity (~33%) Li+ ions contaminat-ed with Na+ (Yoshizuka et al., 2006). Chung et al. (2004) synthesizednano-manganese oxide (Li1.33Mn1.67O4) through a gel process. The ionsieve-type adsorbent containing magnesium after acid treatment was

Recovery (%)/Remarks Product(% purity)

References

l]: 50–1000 mg/L 70% Li – Yoshinaga et al.(1986)

ith 1.8 M Na2CO3

min

Good usable volume cap.of resin TP207:56.4 g/L

Li saltsolution

Bukowsky et al.(1991)

tion on λ-MnO2,

816 m3 seawater);. of Mg(II), Ca(II),II) with SK110 resin

56% yield Li2CO3

(N99.9)Nishihama et al.(2011), Onishi et al.(2010)

11.5–12.5;2CO3 at 100 °C

Recovery of purelithium carbonate

Li2CO3 (99.4) Um and Hirato(2012)

11.3,ium oxalate,

Precipitation as Li2CO3 LiCO3 (99.55) An et al. (2012),Tran et al. (2013)

Mg as hydroxide and LiCl salt feed for Li extr.in a chemical plant

LiCl Clarke (2013)

.5 M NaCl, 10 h ~94% Li;Li sorption: 28.65 mg/gLiFePO4

– Zhao et al. (2013)

Brine

Evaporation concentration

Concentrated brine

Magnesium

hydroxide

Magnesium removal

Filtration /Washing

Calcium removal

Filtration

Lithium

carbonate

Na2CO3

Lithium salt solution

Lime milk

Calcium-lithium-salt solution

Precipitation crystallization

Filtration/Washing

Lithium carbonate

Hydration

RoastingEvaporation

Calcium carbonate

Fig. 1. A flowsheet for Li2CO3 production from brine.


then generated which selectively adsorbed lithium (~30.3 mg/g adsor-bent) from seawater.When a polymeric membrane reservoir containingan inorganic ion-exchange adsorbent inside it with zinc was used, lithi-um recovery from seawater was very effective and kinetically favoredwith adsorption of 33.1 mg Li/g sorbent (Chung et al., 2008). This adsor-bent had excellent lithium adsorption of 89% of 400 mg Li in a day only;the desorption efficiency being 92.88% by dipping in 4 L of 0.5 M HCl so-lution in a day. By using Mg-doped manganese oxide, Chitrakar et al.(2013) also observed very fast adsorption equilibrium (within 24 h) foreffective recovery of lithium from salar brine (Table 7).

RecentlyWajima et al. (2012) prepared HMn2O4 by elution of spinel-type lithium di-manganese-tetra-oxide (LiMn2O4) and examined thekinetics of lithium adsorption. The intermediate, LiMn2O4, was also syn-thesized from LiOH·H2O andMn3O4 by acid treatment. Lithium recoveryfrom seawater reached ~100% at 60 °C using both products.

Aluminium foil immersed in sea water forms a corrosion product onits surfacewhich extracts lithium selectively from seawater at the sametime, at the optimum temperature of ~30 °C (Takeuchi, 1980). Donget al. (2007) prepared an aluminium salt adsorbent using Al(OH)3 andLiOH at pH 5.8, and a molar ratio of 2, and investigated the recovery oflithium from a salt lake bittern by this adsorbent. The adsorbent showedhigh adsorption and uptake of 0.6–0.9 mg Li/g rather than the otheralkali or alkalinemetals of bitterns. Use of hydrated alumina for adsorp-tion of lithiumwas also reported from Egyptian bitterns (Hawash et al.,2010) and salar brines of Argentina (Clarke, 2013).

The Institute of Ocean Energy at Saga University began operating theworld's first — but small — lab aiming for practical lithium productionfrom seawater and succeeded in acquiring about 30 g of lithium chloridefrom 140,000 L of seawater in onemonth. Similarly in early 2010, POSCO

and the Korea Institute of Geoscience and Mineral Resources, Korea hadjoined hands to build a pilot plant for the commercial production oflithium carbonate from sea water based on the adsorption process(USGS, 2009), the outcome of which remains uncertain presently.

3.2.2. Precipitation processSome of the precipitation and other processes used for lithium recov-

ery from seawaters/brines are summarized in Table 8. Among various co-precipitating agents used, aluminium salts show the best performancefor lithium recovery from geothermal water. The appropriate pH forlithium recovery is 10–13 and use of NaAlO2 seems better thanAlCl3. With a high purity NaAlO2 solution as co-precipitating agent,about 98–99% Li recovery was achieved at pH 11.5 from Ca- andSiO2-free geothermal water (Yoshinaga et al., 1986).

Despite being the richest lithium resource (10.2 Mt) Uyuni salarbrine containing a high magnesium concentration (Mg/Li mass ratio~21.2:1) causes difficulties in lithium production. The high magnesiumcontent represents a significant metal value which should be recoveredwith lithium. Consequently, calcium andmagnesiumhad to be removedfrom the brine by using oxalic acid before the production of lithium. TheMg-oxalate producedwas suitable for use as a precursor for the produc-tion of MgO by roasting (Tran et al., 2013). An et al. (2012) developed ahydrometallurgical process to recover lithium from Uyuni salar brinecontaining 15–18 g/LMg and0.7–0.9 g/L Li saturatedwith sodium, chlo-ride and sulfate. In a 2-stage precipitation process, magnesium andsulfate were removed as Mg(OH)2 and gypsum (CaSO4·2H2O) atpH 11.3 by lime in the first stage. Residualmagnesium after lime precip-itation and almost all soluble calcium were then removed by the addi-tion of sodium oxalate. In the second stage 99.6% lithium carbonatewas precipitated at 80–90 °C using sodium carbonate. Residual Li+

from the solution was quantitatively extracted in the presence ofother alkali metals by a mixture of commercial β-diketones (LIX-51)and TBP (Miyai et al., 1988). A 2-stage lime precipitation process totreat seawater was reported by Um and Hirato (2012) to separatelithium from calcium and magnesium, whereas Clarke (2013) men-tioned the use of precipitationmethod to remove the twometals (mag-nesium and calcium) from the salar brine of Argentina to produce LiClwhich was further processed to recover lithium.

3.2.3. Ion exchange/Solvent extraction processFor highmagnesiumand calciumcontaining bitterns or brines solvent

extraction or ion exchange can be used. After selective stripping/elutionlithium can be precipitated out. A combined process consisting of solarevaporation and ion exchange for the extraction of lithiumwas proposedby Steinberg andDang (1975, 1976). TheDowex resinwas used for selec-tive exchange of its H+with the cations present in seawater in the order:K+, Na+, Li+ and Mg2+. Lithium ions were eluted using 0.2–0.5 M HCland eluted LiCl was transferred into an electrolyser to produce lithium.Strelow et al. (1974) separated lithium from sodium, beryllium andmany other elements by eluting lithiumwith 1MHNO3 in 80%methanolfrom a column of AG50W-XS, a sulfonated polystyrene cation-exchangeresin. Samples were loaded onto a 20 cm3 AG50W-XS (200–400 mesh)resin column and then eluted with a mixed acid–methanol solution(1MHNO3 and 80%methanol)which ensured an extremely good separa-tion of lithium from sodium in a single column pass.

Earlier studies showed that organic ion-exchange resins exhibited lowselectivity for lithium ions (Abe and Hayashi, 1984; Alberti andMassucci,1970; Ho et al., 1978). However, Bukowsky et al. (1991) demonstratedthat precipitation followed by ion exchange can be effectively used forseparation and recovery of lithium from a synthetic solution of calciumand magnesium chlorides. Recently Nishihama et al. (2011) appliedSK110 resin (sulfonated type) from a concentrated solution to removedivalent metal ions (Mg2+, Ca2+, Sr2+ and Mn2+) due to theirhigher sorption capacity compared to that of mono-valent ions.The separation of Li+ from the resultant solution with Na+ and K+

was achieved in a packed column of impregnated resin containing

Acetonitrile

CoO

NMP

Water + LiOH

Acetonitrile

50 oC

CoO

Cut up

Spent lithium ion batteries

Steel

PlasticsCu Al

Electrolysis

Decantation

Washing

Solvent evaporation

Sorting of solids

Electrolyte stored

Filtration

Dissolution of binder

Solvent evaporation

NMP

Dry N2 gas

50oC

50-60oC

90 oC

Dissolution of electrolyte

LiOH solution

Fig. 2. Process flowsheet of AEA technology.


1-phenyl-1,3-tetradecanedione (C11phβDK)/tri-n-octylphophine oxide(TOPO). Li2CO3 was precipitated from the concentrated Li+ solution bya (NH4)2CO3 solution with an overall yield of 56% and a purity ofN99.9%. Earlier this group reported selective adsorption of Li+ in aqueouschloride media using a novel synergistic solvent impregnated resin (SIR)containing both 1-phenyl-1,3-tetradecanedione (C11phβDK) and TOPO(Onishi et al., 2010). About 94% Li+ was adsorbed by SIR containing0.66 mmol/g of each extractant at pH 12. Almost 96.9% Li+ was elutedwith a 1.0 mol/L HCl solution with 99.8% purity at a bed volume (BV)of 2.7.

Nelli and Arthur (1970) patented a selective liquid extraction processfor lithium from bitterns of high magnesium content. In the presence ofstrong chloride and HCl, lithium was converted into stable lithium tetrachloroferrate (Cl4FeLi2), which can be extracted by a number of solvents.About 90% Li was extractedwith amixture of 20% TBP and 80% diisobutylketone in 7 counter current stages. Any co-extracted magnesium wasrecycled back to the initial extraction step by washing with water in 4stages. Finally lithium tetra chloroferrate was stripped in 5 countercurrent stages. The stripwaterwith ~2MNaClwas contactedwith amix-ture of solvent (20% D2EHPA and 30% TBP in benzene) in 6 stages toremove FeCl3. The raffinate in this step is the product containing~0.36% Li, 200 mg/L Mg and 20 mg/L Fe. The second solvent is thenstripped of its FeCl3 with about 0.3 parts of water in 6 stages.

Separation of lithium andmagnesium is difficult because they exhibitmany chemical similarities. However, by utilizing the high charge densityof Mg2+which is twice that of Li+ with almost the same ionic radius andits easy hydrated properties, Zhao et al. (2013) reported a separationmethod using LiFePO4/FePO4 as electrodes (Table 8). Lithium exhibitsgood reversibility in LiFePO4/FePO4 structures, and the redoxpeak separa-tion is 0.592 V compared to 1.403 V for Mg2+ indicating its higher polar-ization. At a voltage of 1.0 V in pure lithium solution, the inserted capacityof lithium can reach 41.26mg/g LiFePO4, which is 93.78% of its theoreticalvalue (44 mg). The subsequent extracted capacity of lithium can beattained to 38.93 mg/g LiFePO4, which is 94.3% of its inserted capacity,while the extraction capacity of Mg2+ from a solution containingmagne-sium is only 5.5 mg/g LiFePO4. The lower voltage is found to be beneficialfor separating Mg and Li and the method also works for brine to reducethe Mg/Li ratio to 0.45 from 60.

3.3. Extraction of lithium from secondary resources— lithium ion batteries

Secondary resources like spent batteries have lithium compoundscoupled with valuable metals such as cobalt, nickel, manganese etc.which can be economically viable to recycle,whereas lowvalue elementssuch as iron and phosphorouswill be a greater challenge to create a prof-itable recycling program.World over several companies viz., Toxco, OnTo,Sony, Accurec etc. are currently performing and improving batteryrecycling processes. Though Toxco and Sony were the first to recycleLIBs which they still continue, several others also started processingsuch materials recently (Espinosa et al., 2004; Gaines and Cuenca,2000; Li et al., 2009a). The Toxco process is designed for all types of lith-ium containing wastes. The main product is targeted as appropriate(cobalt/other base metals) and the other component being the lithiumhydroxide. Sony uses incinerationwhich is followed by hydrometallurgi-cal techniques. The industrial processes in vogue/developed for recyclingLIBs are described in brief.

3.3.1. Major industrial processes

3.3.1.1. AEA technology. A patented process developed in the UK aims atthe recovery of all battery materials (Lain, 2002). The casing of the bat-tery is first removed in N2 and cells are cut mechanically to removeother components wound into a spiral. The anode, cathode and separa-tors are shredded (~1 cm2 size) to leach the electrolyte and the solventin acetonitrile at 50 °C. The electrolyte and the solvent are recovered byevaporation of the leach liquor and acetonitrile is recycled back. The

binder-PVDF is dissolved in NMP (N-methyl-2-pyrrolidone) and filteredto separate it fromelectrodematerial (Al, Cu, steel and plastics). TheNMPis evaporated from the filtrate and recycled back. Residue is back-washedand the suspended particulate (LiCoO2 and carbon) is electrolysed withLiOH as electrolyte. LiCoO2 is reduced adjacent to the cathode to formCoO while increasing the concentration of LiOH (Fig. 2) which is recov-ered by decantation. The reaction can be represented as:

e− þ H2O þ LiCoðIIIÞO2→CoðIIÞO þ Li

þ þ 2OH−: ð14Þ

3.3.1.2. Toxco process (Canada). A variety of batteries is processed in cryo-genic conditions (in liquid nitrogen) to reduce the reactivity andmechan-ical and hydrometallurgical approaches are utilized to recover metalvalues (McLaughlin and Adams, 1998). The process flow-diagram isshown in Fig. 3. Large batteries are sheared into three pieces in a causticbath, which neutralizes any acidic components and dissolves the lithiumsalts. The salts precipitated and dewatered in filter press, are used to pro-duce lithiumcarbonate. Besides organics, H2 formedby reactionwith lith-ium, burns off at the surface of the process bath. Sludge fromLIBs is sent torecover cobalt. The large pieces are mechanically treated while the batte-ries are submerged in a process solution of lithium brine, after whichferrous and nonferrous metals are recovered. Li-ion fluff (mixture ofsteel and plastics) can then be separated from the rest of the materials(Cu–Co and slurry). Copper is extracted in the next step and lithiumbrine is treated with Na2CO3 to obtain Li2CO3. Plastics and paper floatingat the top are recovered for disposal or recycling. The filtered carbon

Disposal/ steel

recovery

Li2CO3(s)


Primary metal

production or other

applications

Copper -cobalt

product

Mixing & Holding

tank

Mixing tank

Product filter

Drying

Na2CO3

Water spray

Spent brine

to disposal

Li-ion fluff

Shaker table

Lithium brine

Washing

Filter press

Conveyor

Hammer mill

Slurry

Cobalt filter

cake

Lithium brine

Fig. 3. Toxco's hydrometallurgical recycling process for LIBs.

Acid

Lithium chloride

Al, steel, Cu, plastics

Electronics, plastic casing

Electrolyte, solvent,

conductive salt

Slag former

Residue

Cobalt manganese

alloy


Mechanical pre-treatment

Vacuum thermal treatment

Mechanical crushing

Sieving, magnetic,

air separation

AgglomerationBinder

Reducing meltdown

Leaching

Filtration

< 0.2 mm electrode material

250 oC

Precipitant

Precipitation

Fig. 4. Flow-sheet of Accurec recycling process.

CO2 & inert gas mixture

Precipitation

Li2CO3(s)

Steel, copper, plastics


Mechanical treatment

Physical treatment

Mixed oxides +

inserted Li + Carbon

Leaching of fines

Leaching H2SO4

Purification

Oxidation

Li(aq)

SodaCu and other

impurities

NaClO Li2SO4(aq)

pH > 12

Soluble lithium

Precipitation

Lithium phosphate

Cobalt(III) hydroxide

LiOH(aq)

Fig. 5. Recupyl process for LIBs recycling.


cake from the sludge is un-economical to reuse or even to burn off. Cobaltin the spent LIBs presents higher economic worth.

3.3.1.3. Accurec GmbH (Germany) process. The process involves mechan-ical treatment to separate the electrode material which is treated bypyro-metallurgy to recover the Co–Mn alloy and lithium chloride(Sojka, 1998). Electronics and plastic casings are removed followed byvacuum thermal treatment and pyrolysis to take care of the electrolyteand solvent including conductive salts (Fig. 4). The batteries are thencrushed and by employing sieving, magnetic as well as air separator,aluminium, copper and steel are removed. The remaining electrodematerial is agglomerated with the addition of a binder and pressed tobriquettes, which are smelted (reduction) in a furnace to obtain twofractions, the metallic Co–Mn alloy and lithium containing slag. Byacid leaching of the slag lithium can be extracted as lithium chloride/carbonate.

3.3.1.4. Sony/Sumitomo (Japan) process. In this process untreated LIBs arecalcined at 1000 °C by which the cells open and inflammable compo-nents such as the plastic casing and organic solvent burn off. The residueconsisting of metallic and metal pieces of iron, copper and aluminiumcan be magnetically separated. The remaining fraction is mainly carbonpowder and active cathode material (LiCoO2 and/or LiCoxNi(1−x)O2).The powder is treated at the Sumitomo plant by hydrometallurgicalmethods to recover cobalt; lithium is not targeted. Cobalt oxide ofhigh quality is recovered to use directly in the fabrication of new LIBsand the metallic scraps such as copper and stainless steel can be usedas the by-products. Large batteries can also be handled but needs tobe punctured prior to introduction to the furnace (Gaines and Cuenca,2000; Lupi et al., 2005).

3.3.1.5. Recupyl process (France). This is also called as the Valibat processfor recycling LIBs. Mechanical preparation of the spent batteries is

performed under an inert atmosphere to reduce the reactivity of lithi-um. Plastics, steel and copper are then separated by physical treatment.The fine powder obtained by screening is suspended in stirredwater forsubsequent leaching and hydrolysis step. Filtering the hydrolyzed solu-tion produces an alkaline solution of lithium salt and a suspension of

Acidified

aqueous liquid

Spent lithium batteries

SO2(aq), Cl2(aq)

Recovered chemical

substances

Protective

atmosphereCO2

Feeder &

ShredderCO2

NeutralizationMoist air

Leaching & washingGas Scrubber

Release of protective

atmosphereGases

Residue

Metal fractions

& plastics

Leach liquor

Hydrometallurgical

processing

(SX/Precipitation)


metal oxides and carbon. Lithium is precipitated as Li2CO3 using theCO2 obtained from the mechanical treatment. The suspension of metaloxides is dissolved in H2SO4. Copper is cemented out by the steelshots (Tedjar and Foudraz, 2007). The purified solution is oxidized byNaClO to precipitate cobalt(III) hydroxide (Fig. 5) and cobalt is separat-ed by electrolysis. The remaining lithium in the solution is precipitatedwith CO2 gas.

3.3.1.6. Batrec Industrie AG process (Switzerland). The company Batrecmainly runs a mechanical processing plant for LIBs. In this processbatteries are crushed in CO2 gas atmosphere and the released lithiumis neutralized. With the completion of the neutralization step, theprotective environment is released and subsequently treated in a gasscrubber to reduce the emission during the process. Scrap material isleached and washed in acidified aqueous liquid and leach liquor isfurther processed for the recovery of different chemical substances.Metal containing a solid fraction can be separated from liquid and treat-ed to remove impurities. The flowsheet of the process is presented inFig. 6.

Fig. 6. Batrec Industrie AG process flowsheet.

Table 9Extraction and recovery of metals from spent LIBs in hydrochloric acid system.

Material Leaching conditions Extraction/Sep

Acid concentration(M)

Temp.(°C)

Time(h)

Pulp density(g/L)

Cylindrical-shaped LIBs

4 80 1 100 SX with 0.29 M

Spent LIBs 4 80 1 100 Precipitation: pSynthesis of prco-precipitatio

LIBs 4 80 1 20 Precipitation: ppH 9 for Co by

LIBs 3 M acid + 3.5%H2O2

80 1 50 Precipitation:Na2CO3 at 100

Ash from LIBs 4 90 18 50 Precipitation baddition at pH

3.3.1.7. OnTo Technology (USA). Spent batteries are discharged beforedismantling. Batteries are cleaned and then placed in a high pressureextraction container filled with liquid CO2 and some additives like alkylethers, ammonia etc. The CO2 is transformed into a supercritical fluid byincreasing the pressure and temperature in the container. With anincrease in pressure, the battery casings are breached by fluid. Whenthe desired pressure and temperature are reached the electrolyte extrac-tion can either be carried out by circulating CO2 through the system or bysoaking the batteries in the supercritical fluid. Batteries without function-al potential are subsequently recycled through pulverization and the var-ious materials separated and recovered based on the physical properties.

3.3.1.8. Umicore process (Belgium and Sweden). Also called as the ValÉasprocess, this is a combined pyro-and-hydro-metallurgical process toseparate nickel and cobalt without recovering lithium. The spent batte-ries are smelted in a furnace. Plastic, solvent and graphite are burnt offwhile the metals are reduced and collected in a melt. The process isoperated in Sweden while the molten metal consisting of Ni, Co, Cuand Fe is cooled and processed in Belgium by sulfuric acid leachingand solvent extraction. NiSO4 and CoCl2 are separated from the solutionfrom which cobalt oxide and Ni(OH)2 are obtained.

A disadvantage of pyrometallurgical recycling processes includingthat of Umicore is that lithium recovery is not targeted and for whicha combination of pyrometallurgical and hydrometallurgical processingsteps are necessary. The Umicore process has another drawback oflosing base metals which are slagged off, besides the loss of organicmaterials as well as carbon.

3.3.2. Recent development in recycling of lithium ion batteriesSpent LIBs containing lithium and othermetals aremostly treated

by the hydrometallurgical process which is used at times in combi-nation with pyrometallurgical treatment. This may have integratedpre-treatment steps like pyrolysis ormechanical processing, i.e. crushingand material separation. In order to investigate the extraction of cobalt,nickel, manganese etc. from LIBs by such processes thermodynamic as-pects particularly the stability regions of different phases in aqueous solu-tion under redox conditions may be examined. For this standard Eh–pHdiagrams of Li–H2O, Co–H2O, Mn–H2O and Ni–H2O systems can be re-ferred from literature (Pourbaix, 1966; Schweitzer and Pesterfield, 2010).

Leaching of LIBs are carried out with different acids like HCl(Contestabile et al., 2001; Zhang et al., 1998), H2SO4 (Aktas et al., 2006;Dorella and Mansur, 2007; Kang et al., 2010a,b; Nan et al., 2005; Shinet al., 2005; Swain et al., 2007), HNO3 (Lee and Rhee, 2002, 2003), and afew organic acids like DL-malic acid (Li et al., 2010a), citric acid (Li et al.,2010b) etc. The binder (PVDF) which links the cathode material, LiCoO2

with aluminium foils does not dissolve easily in the organic reagentssuch as fatty hydrocarbon or alcohol at room temperature making theleaching reactionsmore difficult to proceed. The important developments

aration—parameters Highlights (withmerits/demerits)

References

D2EHPA and 0.90 M PC-88A SX: N99.9 Co and 12.6% Li with0.9 M PC-88A.Higher selectivity with PC 88A.

Zhang et al.(1998)

H 6–8 with 4 M NaOH.ecursor material byn.

Co(OH)2 separated easily.NixCoyMnz a precursor to cathodematerial

Contestabileet al. (2001)

H 2 with KMnO4 for Mn(II),DMG and Li2CO3

Purity (%): 96.9 Li, 98.23 Mn,96.94 Co and 97.43 Ni

Wang et al.(2009)

pH 11–12 with NaOH for Co;°C for Li

Recovery (%): 95 Co and 93 Li inprecipitate

Shuva andKurny (2013)

y NaClO at pH 3 and base11

Recovery (%): 100 Co and 99.99 Ni.HCl gives higher leaching efficiency.

Joulié et al.(2014)

Table 10Extraction and recovery of metals from spent LIBs in sulfuric acid system.

Material Leaching conditions Extraction/Separation—parameters

Highlights (withmerits/demerits)

References

Acid concentration Temp.(°C)

Time(h)

Pulp density(g/L)

LIBs waste 2 M acid + 15 vol.%H2O2

75 0.167 50 – Leaching (%): 95 Coand 100 Li.Harmful gas—SOx

generated.

Shin et al.(2005)

Square-shaped LIBs 3 M acid 70 4 200 Precipitation & SX using Acorga M5640and Cyanex 272

Precipitation: 90% CoSX: ~97% Cu by AcorgaM5640, 97% Co by Cyanex272

Nan et al. (2005)

LIBs 6% acid + 1% H2O2 65 33.3 Precipitation with NH3 at pH 5.Co/Li separation by diluted (0.72 M)Cyanex 272

Leaching (%): ~55 Al, 80Co and 95 Li.Extraction (%):~85 Co byCyanex 272.

Dorella andMansur (2007)

LIB wastes 2 M acid + 5% H2O2 75 0.5 100 SX by 0.5 M Cyanex 272, pH 5.35and O/A = 1 in 1-stage

Leaching (%): 93 Coand 94 Li.CoSO4 solution: 99.99%pure.

Swain et al.(2007)

Spent LIBs 4 M acid + H2O2 80 4 Precipitation using ethanol, ethanol/solution = 3:1; 15 min, roomtemperature

Co recovery: 92% asCoSO4 and8% Co as Co(OH)2 byadding Li(OH)2 at pH 10.

Aktas et al.(2006)

Synthetic solution (10 mol/LCo(II) and 20 mol/L Li2SO4)

– – – – SX by Cyanex 272 and DP-8R Separation factor (Co/Li):497 at pH 5

Swain et al.(2010)

Spent LIBs 2 M acid + 6% H2O2 60 1 100 Precipitation at pH 6.5 for Cu, Fe and AlSX by 50% saponified 0.4 M Cyanex 272,pH 6, O/A = 2

Leaching (%): N99 Co,SX: 99.9% Co in 2 stages.Separation factor 750(Co/Li) and Co/Ni atpH 6.

Kang et al.(2010a, 2010b)

Spent LIBs 1 M acid + 30%H2O2

80 2 714 Precipitation by 1 M citric acid, 2 h,65 °C; calcined at 450 °C, 4 h

Max. Co leaching (%): 88.3.Crystalline LiCoO2

synthesized

Li and Zeng(2011)

Spent LIBs 4 M acid + 10%H2O2

85 2 100 SX by 25% P507Precipitation by ammonium oxalate,pH 1.5

SX: 98% Co and removalof 97% Ni and Li.

Chen et al.(2011a, 2011b)

Spent LIBs 3 M + 0.25 MNa2S2O3

90 3 67 – N99% leaching Wang et al.(2012)


on the leaching and separation of metals from the leach solutions aresummarized in Tables 9–11.

3.3.2.1. Metal extraction/recovery from hydrochloric acid leach liquors. Inmost processes reported, optimum extraction of metals from the cath-ode material of spent LIBs was achieved with a 4 M solution of HCl(Table 8). In the presence of a reducing agent such as hydrogen perox-ide, metal leaching was possible even in 3 M HCl (Contestabile et al.,2001; Joulié et al., 2014; Shuva and Kurny, 2013; Zhang et al, 1998).While H2O2 facilitated the dissolution of Co(II) which was reducedfrom Co(III), the dissolution of lithium was also promoted because ofthe presence of the two metals in the same oxide. The separation and

Table 11Extraction and recovery of metals from spent LIBs in other organic/inorganic acid systems.

Material Leachingreagent

Leaching conditions

Time (h) Temp. (°C) Pulp den

LIBs 1 M HNO3 + 1.7 vol.% H2O2 1 75 20

LIBs 1 M HNO3 + 1.0 vol.% H2O2 1 80 20

Spent LIBs 1.5 M DL-malic acid + 2% H2O2 0.67 90 20

Spent LIBs 1.25 M citric + 1% H2O2 0.25 90 20

Cylindrical spent LIBsof mobile phone

1.25 mol/L ascorbic acid 0.33 70 25

recovery of metals from the leach liquors, was carried out either by sol-vent extraction involving PC-88A, Cyanex 272 etc. to produce the puremetal salts or the metals that were selectively precipitated as Co(OH)2and LiCO3 from the leach solutions. Thus, Zhang et al. (1998) used thesteps such as leaching of cathode materials of spent LIBs in HCl, separa-tion of Co/Li by solvent extraction (PC-88A) and recovery of cobalt assulfate and lithium as carbonate. Out of sulfurous acid, hydroxyl aminehydrochloride and hydrochloric acid, HClwas found to be themost suit-able lixiviant for economic reasons. The leaching efficiency of N99% ofCo and Li in a 4 M HCl solution was achieved. The complete extractionof cobalt from the leach liquor by 0.90 M PC-88A in 1 stage at pH 6.7and O/A of 0.85:1, scrubbing of lithium by 30 g/L Co at O/A 10:1 and

Highlights(with merits/demerits)

References

sity (g/L)

Leaching (%): 95 Co and Li.Harmful gas—NOx generated.

Lee and Rhee (2002, 2003)

Leaching (%): 100 Co and Li.Harmful gas—NOx generated.

Li et al. (2011)

Leaching (%): 90 Co and 100 Li.Acid can be recycled and reused.

Li et al. (2010a)

Leaching (%): N90 Co and 100 Li.Simple and environmental friendly process.

Li et al. (2010b)

Leaching (%): 94.8 Co and 98.5 Li.Reducibility of ascorbic acid displaced H2O2.

Li et al. (2012)


stripping of cobalt from the loaded organic with 2 M H2SO4 at O/A 5:1,produced high purity cobalt sulfate. About 80% lithium was recoveredas Li2CO3 by precipitation with saturated soda solution at 100 °C.The leaching reaction of the waste LiCoO2 with HCl solution can berepresented as:

4LiCoO2ðsÞ þ 12HClðaqÞ→4LiClðaqÞ þ 4CoCl2ðaqÞ þ 6H2O þ O2ðgÞ: ð15Þ

A laboratory scale process involving leaching of LiCoO2 inHCl and pre-cipitation of cobalt hydroxidewas developedbyContestabile et al. (2001).The activematerial of spent LIBswas dissolved inN-methylpyrrolidone at100 °C to separate aluminium and copper foil. From the leach liquor, car-bon powderwas removed by filtration and Co(OH)2 was precipitated outfrom the solution at pH 6–8. In another study over 95% dissolution of Ni,Co and Mn was reported when the battery material of 120 μm size wasleached at a higher acid concentration (6 M HCl) and lower temperature(60 °C) in the presence of H2O2 (Li et al., 2009b). Cobalt from the solutionwas cemented out by iron powder and iron was removed as goethite.Finally chlorides of Ni, Mn and Co were added in the purified solutionto prepare the precursor of the cathode material (NixCoyMnz) directlythrough co-precipitation with ammonium bicarbonate. The leaching ofmetals from the cathode material in 4 M HCl at 80 °C with a metal re-covery of 99%was also reported byWang et al. (2009). Manganese inthe leach liquor was precipitated at a molar ratio of Mn2+ to KMnO4

of 2 and pH 2 followed by selective precipitation of nickel withdimethylglyoxime (DMG) at pH 9. Cobalt was recovered as hydrox-ide at pH 11 and Li2CO3 was then precipitated; purity of the metalsbeing 96.97% Li, 98.23% Mn, 96.94% Co and 97.43% Ni. A recent studyby Shuva and Kurny (2013) demonstrated the reductive dissolution ofcathode powder in 3 M HCl in the presence of 3.5% H2O2. Over 95%cobalt was precipitated as hydroxide at pH 11–12, leaving lithium(93%) in the leach solution.

Processing of battery ash obtained from the pyrolysis of spent LIBs byHCl leachingwas also attempted. Lin et al. (2003) patented a pyrometal-lurgical process of waste LIBs combined with hydrometallurgical pro-cessing. The ash with metal and metal oxides was dissolved in 3–6 MHCl containing NaCl. Copper and cobalt were separated out using mem-brane electrolysis. Then Fe(OH)3 and Al(OH)3 were recovered at pH 5–7followed by the precipitation of lithium carbonate. Recently Joulié et al.(2014) reported the high leaching efficiency (~100%) of Li, Ni, Co andAl from the Li–Ni–Co–Al oxide ash of spent LIBs by 4 M HCl at 90 °Cwith chloride ions promoting the dissolution. Co(II) in the leach liquorwas oxidized to Co(III) with NaClO and recovered as Co2O3·3H2O byselective precipitation at pH 3 (Eqs. (16) and (17)). Nickel hydroxidewas then precipitated at pH 11.

2Co2þ þ ClO

− þ 2H3Oþ⇔2Co

3þ þ Cl− þ 3H2O ð16Þ

2Co3þ þ 6OH

−→Co2O3⋅3H2O: ð17Þ

3.3.2.2. Metal extraction/recovery from sulfuric acid leach liquors. In mostleaching processes with sulfuric acid, hydrogen peroxide was used asa reductant (Castillo et al., 2002; Dorella and Mansur, 2007; Jha et al.,2013a,b; Kang et al., 2010a, 2010b; Shin et al, 2005). In some cases alkalileaching followed by acid leaching was considered to remove alumini-um. Metals from the leach solutions were separated and recovered bysolvent extraction using PC-88A/P507, Cyanex 272 and Acorga M5640and precipitation processes very similar to that of the HCl system. Thefine sized electrodes were initially contacted with N-methyl pyrrol-idone (NMP) to dissolve the binder and separate active material(LiCoO2) from Al and Cu foils (Castillo et al., 2002). The LiCoO2 powderwas leached in 4 M H2SO4 at 80 °C to dissolve Co and Li, and Co(OH)2was recovered from the leachate by the addition of sodium hydroxide.

Shin et al. (2005) reported almost quantitative leaching of cobalt and lith-iumwith 2 MH2SO4 in the presence of a high amount of H2O2 (15 vol.%)at 75 °C and 50 g/L pulp density (Table 10). Reductive leaching of themechanically treated LIBs in 6 (v/v)%H2SO4 and1%H2O2 solution resultedin a relatively lower recovery of metals (~55% Al, 80% Co and 95% Li)(Dorella and Mansur, 2007). Kang et al. (2010a, 2010b) also reportedthe reductive leaching of LIBs with H2SO4. The reactions of LiCoO2 withH2SO4 and in the presence of H2O2 are shown below:

4LiCoO2ðsÞ þ 6H2SO4→4CoSO4 þ 2Li2SO4 þ 6H2O þ O2ðgÞ ð18Þ

2LiCoO2ðsÞ þ 3H2SO4 þ H2O2ðaqÞ→4CoSO4 þ Li2SO4 þ 4H2O

þ O2ðgÞ: ð19Þ

The leaching efficiency of cobalt depends on the reductant concentration.Among other reducing agents, Na2S2O3 helped in the leaching of N99%of the metals (Co and Li) in 2 M sulfuric acid at 90 °C in 3 h (Wanget al., 2012).

Safe dismantling procedures of spent lithium ion batteries haveoften been described in the literature (Nan et al., 2005; Tanii et al.,2003). Zhu et al. (2011) applied a mechanical separation process torecover copper from these batteries. The anodes from the batterieswere separated by mechanical treatment, pulverization and sieving.Almost 92% of copper in anode particles was recovered by a gas-fluidized bed separator.

In the alkali–acid leachingprocess, the cathodewasfirst treatedwith10% (w/w) NaOH at 30 °C to dissolve Al, followed by reductive leachingof ~97% Co and 100% Li with H2SO4 and H2O2 (Ferreira et al., 2009; Nanet al., 2005). Acorga M5640 and Cyanex 272 were used to selectivelyextract and recover 98% Cu and 97% Co, respectively from the solutions.About 90% Co was recovered as oxalate with b0.5% impurities. LiCoO2

positive electrode material with a good electrochemical performancewas synthesized by using the recovered compounds.

From the sulfate leach liquor of spent LIBs, 96% copper was recov-ered as CuSO4·3H2O with ethanol at a volume ratio of 3:1. Cobalt wasrecovered in two steps. During the first step, 92% of the cobalt wasrecovered as CoSO4 by the use of ethanol at a volume ratio of 3:1. Theremaining Co in the second step was recovered as Co(OH)2 by the addi-tion of Li(OH)2 at pH 10 (Aktas et al., 2006). Lithium,which remained inthe solution, was then recovered to the extent of 90% as Li2SO4 by theaddition of ethanol (3:1 volume ratio). Aluminium was recovered asAl(OH)3 with 99% recovery efficiency. It was shown that metals couldbe precipitated/separately by the ethanol/sulfate precipitation tech-nique depending on their concentrations present in the solution.

Thewaste cathodic activematerial generated during themanufactur-ing of LIBs was also leached in H2SO4 in the presence of H2O2 (Swainet al., 2007). During the separation of Co/Li in a 2-step SX process with1.5MCyanex 272 at O/A 1.6, about 85% Cowas recovered. The remainingcobalt was extracted in 0.5 M Cyanex 272 at O/A 1 and pH 5.35. Thepurity of cobalt sulfate in solution was found to be 99.99%. Earlier,Swain et al. (2006) reported the highest separation factor (Co/Li) of 62during extraction with saponified Cyanex 272 from a synthetic solutionat pH 6.9.

The mechanism by which a metal ion is extracted from an aqueoussolution using a partially saponified cation exchange extractant is asfollows (Ritcey and Ashbrook, 1984):

M2þaq þ A

−org þ 2ðHAÞ2org↔MA2⋅3HAorg þ H

þaq: ð20Þ

Cobalt is extracted as [CoA2·3HA]org with 65% Na-Cyanex 272. Cobalt canbe completely stripped from the loaded organic with 0.01 M H2SO4 toproduce cobalt sulfate of N99% purity.

By acid leaching of waste cathodic material and SXwith Cyanex 272,Swain et al. (2008) produced a pure cobalt sulfate solution (99.99%). Inanother study quantitative separation of Co(II)/Li was reported using


the supported liquid membrane (SLM) with a mixed extractant contain-ing Cyanex 272 and DP-8R as the mobile carrier (Swain et al., 2010).Very recently, leaching of lithium and cobalt from cathodic material ofwaste mobile phone batteries in sulfuric acid in the presence of H2O2

and separation of metals by Cyanex 272 were reported (Jha et al.,2013a, 2013b). Chen et al. (2011a) recovered cobalt oxalate from thespent LIBs by using an alkali–acid leach process. After roasting the spentLIBs at 700–800 °C to burn off carbon and the binder, and leaching withNaOH to remove Al, and leaching with H2SO4 in the presence of 10 (v)%H2O2 could recover 95% Co and 96% Li. Iron removal (99.99%) as jarositewith a loss of b1% Co was achieved in the pH range 3–3.5 at 95 °C. Theprecipitation reaction is shown below:

Fe2ðSO4Þ3 þ 12H2O þ Na2SO4→Na2Fe6ðSO4Þ4ðOHÞ12 þ 6H2SO4: ð21Þ

Complete removal of manganese with ammonium persulfate at pH 4 and70 °C is as follows:

Mn2þ þ ðNH4Þ2S2O8 þ 2H2O→MnO2 þ ðNH4Þ2SO4 þ H2SO4

þ 2Hþ: ð22Þ

Copper precipitation (N98.5%) with NaOH was followed by solventextraction to recover Co(II) while removing 97% Ni and Li (Zhu et al.,2012). About 95% Co(II) was extracted selectively from the purifiedsolution with saponified 25 wt.% P507 (2-ethylhexyl phosphonic acidmono-2-ethylhexyl ester) at pH3.5 and strippedwith 3MH2SO4. Cobaltoxalate was produced from the strip liquor with a purity of 99.9%.CoC2O4·2H2O and Li2CO3 were also produced by a combination of acidleaching, SX and precipitation from spent mobile phone batteries (Jianet al., 2012). From the leach solution of the used batteries, Co(OH)2was precipitated and was converted to Co3O4 by heating (Yamaji et al.,2011) as per reactions (23) and (24). Lithiumwas recovered as carbonateby adsorption with 2% MnO2 (11.7 mg Li/g).

CoSO4 þ 2NaOH→CoðOHÞ2 þ Na2SO4 ð23Þ

3Co OHð Þ2 þ12O2→Co3O4 þ 3H2O: ð24Þ

The advantage of vacuum pyrolysis to prevent the escape of toxicgases and lower the decomposition temperature of organics whileprotecting oxidization of metals in LIBs, was applied by Sun and Qiu

Acorga M5640

Sulfuric acid Leaching of spent LIBs

Al(III), Cu(II), Co(III), Li(I)

Solvent Extraction

Al(III), Co(II), Li(I)Cu(II)

pH 1.5-2.0

pH 2.5-3.0

Solvent Extraction

Co(II), Li(I)

PC-88A

Al(III)

Solvent ExtractionPC-88A + TOA

Li(I)Co(II)

pH 5.5-6.0

Fig. 7. SX separation of Al, Cu, Co and Li from sulfate media (Suzuki et al., 2012).

(2011). The cathode material was treated in a vacuum furnace at600 °C for 30 min with the residual gas pressure of 1.0 kPa. Over 99%Co and Li were recovered from peeled Co–Li oxide by leaching with2 M H2SO4 at 80 °C. Paulino et al. (2008) examined two approaches forrecycling spent Li/MnO2 and LIBs. In the first process ~90% Li was recov-ered by calcination at 500 °C followed by leaching with H2SO4 and H2O2

at 90–100 °C, and solvent extraction. The second process involved fusionwith KHSO4 at 500 °C andwater leaching with H2O2 at 90 °C. Cobalt andmanganese were precipitated at pH N9 followed by precipitation of LiFby KF solution.

Extraction of Co and Li from a material of a large scale mechanicalpre-treatment and recycling plant in Northern Italy is described byGranata et al. (2012). The powder is leached in H2SO4 in the presenceof 50% excess of a reducing agent, glucose — a waste product of thefood factory. Iron, aluminium and copper are partially precipitatedas hydroxides at pH 5.0. Using solvent extraction high purity cobaltcarbonate (47% Co) is obtained by precipitation, whereas without sol-vent extraction the product containing 36–37% (w/w) Co is obtained.Lithium is recovered by crystallization (yield 80%) with 98% purity.The process with solvent extraction shows economical outputs (grossmargin and payback time) than the one without solvent extraction.

Suzuki et al. (2012) developed a process as detailed in Fig. 7. AcorgaM5640 extracted copperwithin a pH range of 1.5–2.0 leaving Al, Co andLi in the raffinate. Aluminium is then selectively extracted by PC-88A inthe pH range 2.5–3.0. Cobalt(II) and lithium(I) are separated by PC-88A/TOA rather than Acorga M5640 due to its higher stripping efficiency(N98%), although Acorga M5640 provides higher cobalt selectivity.

Synergistic extraction and separation of Co(II) and Mn(II) withLi(I) from simulated sulfuric acid leaches of waste cathodic materialsusing a mixture of Cyanex 272 and PC-88A in N-heptane have beeninvestigated by Zhao et al. (2011). A mixed extractant system was alsoutilized to treat the leach solutions of spent LIBs (Pranolo et al., 2010).In the first stage Fe(III), Al(III) and Cu(II) were extracted using Ionquest801 and Acorga M5640, and the raffinate containing cobalt, nickel andlithium was treated with 15% (v/v) Cyanex 272 to separate out Co.More than 90% of Co could be separated out at pH 5.5–6.0 and A/O1:2. An ion-exchange resin such as Dowex M4195 was then used toseparate Ni(II)/Li(I). The addition of 2% (v/v) Acorga M5640 to 7% (v/v)Ionquest 801 generated a significant pH isotherm shift for Cu whichresulted in a ΔpH50 of 3.45.

Apart from the hydrometallurgical extraction of valuable metalsfrom spent LIBs, synthesis of LiCoO2 was also reported (Li and Zeng,2011). The solvent N-methyl-2-pyrrolidone was used to dissolve thePVDF agglomerate from Co–Li membrane to separate the aluminiumfoil. About 88% Co was recovered during the leaching with sulfuricacid at pH 0.5 and 80 °C. Then, 1 M citric acid solution was added at65 °C to prepare a gelatinous precursor. LiCoO2 was obtained with thegel precursor calcined in a crucible at 450 °C for 4 h.

3.3.2.3. Metal extraction/recovery from nitric acid leach liquors.Mechanical,thermal, hydrometallurgical and sol–gel steps were applied to recoverCo/Li from spent LIBs and synthesize LiCoO2 as a cathode activematerial(Lee and Rhee, 2002, 2003). Reductive leaching in 1 M HNO3 with1.7 vol.% H2O2 at 75 °C extracted ~95% Li and 95% Co (Table 11). Themolar ratio of lithium to cobalt in the leach liquor was adjusted to 1.1by adding a fresh LiNO3 solution. Then, 1 M citric acid solution wasadded to prepare a gelatinous precursor which was calcined at 950 °Cfor 24 h to produce purely crystalline LiCoO2. Like in other studies, theaddition of hydrogen peroxide increased the leaching efficiency ofmetals (Lee and Rhee, 2003). The reductive leaching of almost 100%Co and Li from the spent batteries in nitric acid and H2O2 was achievedat 80 °C by Li et al. (2011). In the nitric acid leaching system pollutiondue to harmful NOx gases is a major problem and needs to be addressed.As regards the leaching kinetics, lithium dissolution followed a shrinkingcore model. The mechanism of the dissolution of LiCoO2 was controlledby a surface chemical reaction with an apparent activation energy of


52.32 and 47.72 kJ/mol for Co and Li, respectively (Lee and Rhee, 2003).The leaching reaction with HNO3 solution can be represented as:

2LiCoO2 sð Þ þ 3HNO3 aqð Þ→Li2NO3 aqð Þ þ 2Co NO3ð Þ2 aqð Þ þ 3H2O

þ 12O2 gð Þ: ð25Þ

Recycling of Li-ion and polymer batteries while producingLiCoxNi(1−x)O2 as a cathode material was investigated by Lupi andPasquali (2003). The process consists of cathodic paste leaching, Co–Niseparation by SX and recovery of Ni by electrowinning. The separationof Ni/Co was performed by solvent extraction using saponified 0.5 MCyanex 272. Nickel was electrowon at a current density of 250 A/m2,50 °C and pH 3–3.2 with an electrolyte of 50 g/L Ni and 20 g/L H3BO3.Elaborating further, Lupi et al. (2005) reported the current efficiencyand energy consumption of 87% and 2.96 kWh/kg for nickel under theabove conditions as compared to the figures of 96% and 2.8 kWh/kg,respectively for cobalt at the same current density and temperature,but at pH 4–4.2 from a solution containing manganese and (NH4)2SO4.

3.3.2.4. Metal extraction using organic acids/reagents. For the sustainablemanagement of the secondary resource such as LIBs, organic acidssuch as DL-malic acid, citric acid etc. which have mild acidity, are sug-gested for the leaching of metals (Table 11). Li et al. (2010a) reportedthat DL-malic acid can dissolve lithium and cobalt of LIBs fairly rapidlyunder aerobic and anaerobic conditions as compared to the mineralacids like HCl, HNO3 and H2SO4, and the waste solutions can be treatedeasily. Almost 100% Li and N90% Co were leached out with 1.5 M malicacid and 2.0% H2O2 at 90 °C in 40 min. Leaching of lithium and cobaltin citric acid was also reported after separating the anode and cathodematerial by the treatment of NMP (Li et al., 2010b). Nearly 100% Liand N90% Cowere extracted in 1.25M citric acid and 1.0%H2O2 at 90 °C.

Recently ultrasonic assisted leaching of cobalt and lithium fromspent LIBs in the presence of ascorbic acid was investigated (Li et al.,2012). Leaching efficiencies of as high as 94.8% for Co and 98.5% for Liwere achieved with 1.25 M ascorbic acid solution at 70 °C (Table 11).Sun and Qiu (2012) used oxalic acid as both leachant and precipitantto separate and recover cobalt and lithium from the spent LIBs. Cathodematerial consisting of LiCoO2 and CoO from the dismantled batteries,was peeled off from the aluminium foils after vacuum pyrolysis at600 °C. Leaching was performed using 1 M oxalate at 80 °C with thereaction efficiency of N98% of LiCoO2while separating cobalt and lithium.The reaction with oxalate for leaching and precipitation proceeds as:

3H2C2O4 þ LiCoO2ðsÞ→LiHC2O4 þ CoðHC2O4Þ2 þ 2H2O þ 2CO2ðgÞ ð26Þ

4H2C2O4 þ 2LiCoO2ðsÞ→Li2C2O4 þ 2CoC2O4ðsÞ þ 4H2O þ 2CO2ðgÞ: ð27Þ

During the oxalate leaching Co3+ was reduced to Co2+ which wasdissolved and precipitated as cobalt(II) oxalate (Sun and Qiu, 2012).The reduction of Co3+ to Co2+ is also believed to proceed by the reactionof CO2 radicals generated from oxalic acid (Hoffman and Simic, 1973).

The electrochemical performance of nano-Co3O4 anode materialprepared from the spent LIBs was evaluated by Hu et al. (2013). Fromthe leach liquor obtained from the alkali and acid process, Al(OH)3,MnOOH, Cu(OH)2 and Ni(OH)2 were removed at a pH N5 which wasfollowed by the precipitation of CoC2O4 at pH 2 by adding a saturatedsolution of (NH4)2C2O4. The product CoC2O4 was used to synthesizenano-Co3O4 by the sol–gel method.

The roasting of batteries under reduced pressure at 650 °C wasreported by Kondás et al. (2006), which was followed by the leachingof Li2CO3 at ambient temperature and crystallization of pure Li2CO3. Aprocess for the recycling and synthesis of LiCoO2 from the incisorsbound of Li-ion batteries was developed by Liu et al. (2006). Firstly,

LiCoO2 was separated from aluminium foil with dimethyl acetamide(DMAC). Polyvinylidene fluoride and carbon powders in the activematerial were then eliminated by roasting at 450 °C for 2 h and 600 °Cfor 5 h, respectively. Finally LiCoO2 was obtained by adding a certainamount of Li2CO3 in the recycled LiCoO2 and calcining it at 850 °C for12 h.

4. Conclusions

Lithium is one of the rare metals with a variety of applications anddemand for lithium is expected to increase with the ever increasinguse of electrical and electronic devices/hybrid electric vehicles. A fewestablished technologies are in vogue to produce lithium in the desiredform from its primary resources like its minerals and brine, whereaslimited exploitation of lithium resources from seawater and bitternscalls for their intensified tapping. Extraction of lithium from itsmineralsand clays is fraught with high mining costs and involves high energy,while extraction from brine and bitterns/seawater needs a long timefor evaporation. Hence these processes need to be adequately modifiedto yield efficiency and better economic returns.

Extraction processes from secondary resources like batteries dependon the chemistry of the battery material. Most processes involve dis-mantling of LIBs, separation of cathode and anode materials, leachingof valuable metals like Co, Li, Ni, Mn etc. from the cathode material indifferent mineral acids, and separation and recovery of metals fromthe solutions by solvent extraction/IX/precipitation. At present no lithi-um extraction is industrially practiced from LIBs. Therefore, a sufficientscope exists not only to reduce the process steps followed currently butalso to improve the efficiency of metal extraction and separation, in-cluding lithium recovery. Points thatmay be considered include processintensification to save energy and improve the kinetics of leaching,while addressing the problem of selectivity and examining the use ofunexplored/synergistic solvents with an ultra highmetal loading capac-ity to cut down the process steps. There is a need to develop appropriatetechnology which can address the limitation of current processes forextraction of all valuable metals from its primary as well as secondaryresources.

Acknowledgements

The authors are thankful to the Director, CSIR-National MetallurgicalLaboratory (NML), Jamshedpur, India for giving permission to publishthe paper.

References

Abe, Y., 2010. Rare metal series — current status of lithium resources. JOGMEC MineralResources Report.

Abe, M., Hayashi, K., 1984. Synthetic inorganic ion-exchange materials. XXXIV.Selective separation of lithium from seawater by tin(IV) antimonate cationexchanger. Hydrometallurgy 12, 83–93.

Aktas, S., Fray, D.J., Burheim, O., Fenstad, J., Acma, E., 2006. Recovery of metallic valuesfrom spent Li ion secondary batteries. Miner. Process. Ext. Metall. (Trans. Inst. Min.Metall. C) 115 (2), 95–100.

Alberti, G.M., Massucci, M.A., 1970. Crystalline insoluble salts of tetravalent metals. IX.Thorium arsenate, a new inorganic ion-exchanger, specific for lithium. J. Inorg.Nucl. Chem. 32, 1719–1727.

Amer, A.M., 2008. The hydrometallurgical extraction of lithium from Egyptianmontmorillonite-type clay. J. Met. 60 (10), 55–57.

An, J.W., Kang, D.J., Tran, K.T., Kim, M.J., Lim, T., Tran, T., 2012. Recovery of lithium fromUyuni salar brine. Hydrometallurgy 117–118, 64–70.

Australia's Mineral Resource Assessment (AMRA), 2013. http://www.ga.gov.au/corporate_data/77481/77481_AMRA_v2.pdf.

Averill, W.A., Olson, D.L., 1978. A review of extractive processes for lithium from ores andbrines. Energy 3, 305–313.

Bach, R.O., Wasson, J.R., 1981. Lithium and lithium compounds, 3rd edition Ch. in Kirk-Othmer: Encyclopedia of Chemical Technology vol. 14. Wiley, pp. 448–476.

Banerjee, D.C., Krishna, K.V.G., Murthy, G.V.G.K., Srivastava, S.K., Sinha, R.P., 1994.Occurrence of spodumene in the rare metal bearing pegmatites of Marlagalla–Allapatna area, Mandya District, Karnataka, India. J. Geol. Soc. India 44,127–139.

http://refhub.elsevier.com/S0304-386X(14)00227-8/rf0665















http://www.ga.gov.au/corporate_data/77481/77481_AMRA_v2.pdf

http://www.ga.gov.au/corporate_data/77481/77481_AMRA_v2.pdf









Barbosa, L.I., Valente, G., Orosco, R.P., González, J.A., 2014. Lithium extraction from β-spodumene through chlorination with chlorine gas. Miner. Eng. 56, 29–34.

Bernardes, A.M., Espinosa, D.C.R., Tenório, J.A.S., 2004. Recycling of batteries: a review ofcurrent processes and technologies. J. Power Sources 130 (2004), 291–298.

Brown, J.C., Dey, A.K., 1955. India's Mineral Wealth, 3rd edition. Geoffrey Cumberlege,Oxford University Press, Bombay, p. 271.

Bukowsky, H., Uhlemann, E., Steinborn, D., 1991. The recovery of pure lithium chloride from“brines” containing higher contents of calcium chloride and magnesium chloride.Hydrometallurgy 27 (3), 317–325.

Castillo, S., Ansart, F., Laberty, R., Portal, J., 2002. Advances of recovering of spent lithiumbattery compounds. J. Power Sources 112, 247–254.

Chen, L., Tang, X., Zhang, Y., Li, L., Zeng, Z., Zhang, Y., 2011a. Process for the recovery ofcobalt oxalate from spent lithium-ion batteries. Hydrometallurgy 108, 80–86.

Chen, Y., Tian, Q., Chen, B., Shi, X., Liao, T., 2011b. Preparation of lithium carbonatefrom spodumene by a sodium carbonate autoclave process. Hydrometallurgy109, 43–46.

Chitrakar, R., Kanoh, H., Miyai, Y., Ooi, K., 2001. Recovery of lithium from seawater usingmanganese oxide adsorbent (H1.6Mn1.6O4) derived from Li1.6Mn1.6O4. Ind. Eng. Chem.Res. 40, 2054–2058.

Chitrakar, R., Makita, Y., Ooi, K., Sonoda, A., 2013. Magnesium-doped manganese oxidewith lithium ion-sieve property: lithium adsorption from salt lake brine. Bull. Chem.Soc. Jpn. 86 (7), 850–855.

Chung, K.S., Lee, J.C., Kim, E.J., Lee, K.C., Kim, Y.S., Ooi, K., 2004. Recovery of lithium fromseawater using nano-manganese oxide adsorbents prepared by gel process. Mater.Sci. Forum 449–452, 277–280.

Chung, K.S., Lee, J.C., Kim, W.K., Kim, S.B., Cho, K.Y., 2008. Inorganic adsorbent containingpolymeric membrane reservoir for the recovery of lithium from seawater. J. Membr.Sci. 503–508.

Clarke, G.M., 2013. Lithium-ion batteries: raw material considerations. Am. Inst. Chem.Eng. 44–52.

Clarke, G.M., Harben, P.W., 2009. Lithium Availability Wall Map (LAWM), (Self published& distributed).

Contestabile, M., Panero, S., Scrosati, B., 2001. A laboratory-scale lithium ion batteryrecycling process. J. Power Sources 92, 65–69.

Crocker, L., Lien, R.H., 1987. Lithium and its recovery from low-grade Nevada clays. Bul-letin 691. US Bureau of Mines, Department of Interior, (http://digicoll.manoa.hawaii.edu/techreports/PDF/USBM-691.pdf).

Dong, Q., Li, Y., Piao, X., Zhu, S., 2007. Recovery of lithium from salt lake bittern usingaluminium salt adsorbent. Chin. J. Rare Met. 3, 357–361.

Dorella, G., Mansur, M.B., 2007. A study of the separation of cobalt from spent Li-ionbattery residues. J. Power Sources 170, 210–215.

Espinosa, D.C.R., Bernardes, A.M., Tenório, J.A.S., 2004. An overview on the current processesfor the recycling of batteries. J. Power Sources 135, 311–319.

Evans, R.K., 2008. An Abundance of Lithium. http://www.che.ncsu.edu/ILEET/phevs/lithium-availability/An_Abundance_of_Lithium.pdf.

Evans, R.K., 2010a. The lithium–brine reserve conundrum. North. Miner. 96 (35), 2–6.Evans, R.K., 2010b. Lithium's future supply, demand. North. Miner. 96 (35), 11–13.Fergus, J.W., 2010. Recent development in cathodematerials for lithium ion batteries.

J. Power Sources 195, 939–954.Ferreira, D.A., Prados, L.M., Majuste, D., Mansur, M.B., 2009. Hydrometallurgical separation

of aluminium, cobalt, copper and lithium from spent Li-ion batteries. J. Power Sources187, 238–246.

Ferrell, J.E., 1985. Lithium. Mineral Facts and Problems (1985). U.S. Dept. of the Interior,Bureau of Mines, Washington, D.C., pp. 461–470.

Gaines, L., Cuenca, R., 2000. Costs of lithium-ion batteries for vehicles. United StatesDepartment of Energy, Center for Transportation Research: Energy Systems Division,Argonne National Laboratory, Argonne.

Garrett, D.E., 2004. Handbook of Lithium and Natural Calcium Chloride: TheirDeposits, Processing, Uses and Properties, 1st ed. Elsevier, Amsterdam, TheNetherlands 0-12-276152-9.

Granata, G., Moscardini, E., Pagnanelli, F., Trabucco, F., Toro, L., 2012. Product recoveryfrom Li-ion battery wastes coming from an industrial pre-treatment plant:lab scale tests and process simulations. J. Power Sources 206, 393–401.

Gruber, P.M.P., 2010. Global Lithium AvailabilityThesis—Master in Science University ofMichigan, USA.

Gruber, P.W., Medina, P., Keoleian, G.A., Kesler, S.E., Everson, M.P., Wallington, T.J., 2011.Global lithium availability. J. Ind. Ecol. 15 (5), 760–775.

Hawash, S., Abd, E.K.E., El, D.G., 2010.Methodology for selective adsorption of lithium ionsonto polymeric aluminium (III) hydroxide. J. Am. Sci. 6 (11), 301–309.

Ho, P.C., Nelson, F., Kraus, K.A., 1978. Adsorption on inorganic materials: VII. Hydrous tinoxide and SnO2-filled carbon. J. Chromatogr. A 147, 263–269.

Hoffman, M.Z., Simic, M., 1973. Rate of reduction of cobalt(III) complexes by car-bon dioxide (CO2•

−) radicals in aqueous solution. Inorg. Chem. 12 (10),2471–2472.

Holdren, J.P., 1971. Adequacy of lithium supplies as a fusion energy source. LawrenceLivermore Laboratory, UCID-15953.

Hu, C., Guo, J., Wen, J., Peng, Y., 2013. Preparation and electrochemical performance of Nano-Co3O4 anode materials from spent Li-ion batteries for lithium ion batteries. J. Mater. Sci.Technol. 29 (3), 215–220.

Industrial Minerals & Rocks: Commodities, Markets, and Uses 2006. 7th Edition, Ed. KogelJ.E., Trivedi N.C., Barker J.M., Krukowski S.T. USA. ISBN-13: 978-0-87335-233-8. ISBN-10: 0-87335-233-5.

Jandová, J., Vu, H.N., Belková, T., Dvořák, T., Kondás, J., 2009. Obtaining Li2CO3 fromzinnwaldite wastes. Ceramics-Silikáty 53 (2), 108–112.

Jandová, J., Dvořák, P., Vu, Hong N., 2010. Processing of zinnwaldite waste to obtainLi2CO3. Hydrometallurgy 103 (1–4), 12–18.

Jha, M.K., Kumari, A., Jha, A.K., Kumar, V., Hait, J., Pandey, B.D., 2013a. Recovery of lithiumand cobalt from waste lithium ion batteries of mobile phone. Waste Manag. 33 (9),1890–1897.

Jha, A.K., Jha, M.K., Kumari, A., Sahu, S.K., Kumar, V., Pandey, B.D., 2013b. Selective separa-tion and recovery of cobalt from leach liquor of discarded Li-ion batteries usingthiophosphinic extractant. Sep. Purif. Technol. 104, 160–166.

Jian, G., Guo, J., Wang, X., Sun, C., Zhou, Z., Yu, L., Kong, F., Qiu, J., 2012. Study on separationof cobalt and lithium salts from waste mobile-phone batteries. Proced. Environ. Sci.16, 495–499.

Joulié, M., Laucournet, R., Billy, E., 2014. Hydrometallurgical process for the recovery ofhigh value metals from spent lithium nickel cobalt aluminium oxide based lithium-ion batteries. J. Power Sources 247, 551–555.

Kang, J., Senanayake, G., Sohn, J., Shin, S.M., 2010a. Recovery of cobalt sulphate from spentlithium ion batteries by reductive leaching and solvent extraction with Cyanex 272.Hydrometallurgy 100, 168–171.

Kang, J.G., Sohn, J.S., Chang, H.K., Senanayake, G., Shin, S.M., 2010b. Preparation of cobalt oxidefrom concentrated cathode material of lithium ion batteries by hydrometallurgicalmethods. Adv. Powder Technol. 21, 175–179.

Kang, D.H.P., Chen, M., Ogunseitan, O.A., 2013. Potential environmental and human healthimpacts of rechargeable lithium batteries in electronic waste. Environ. Sci. Technol.47, 5495–5503.

Karstetter B.R., 1971. Glass-ceramic article and method. US patent 3,585,054. June 15.Kesler, S.E., Gruber, P.W., Medina, P.A., Keoleian, G.A., Everson, M.P., Wallington, T.J., 2012.

Global lithium resources: relative importance of pegmatite, brine and other deposits.Ore Geol. Rev. 48, 55–69.

Kipouros, G.J., Sadoway, D.R., 1998. Towards new technologies for the production oflithium. J. Met. 50 (5), 24–26.

Kitajou, A., Suzuki, T., Nishihama, S., Yoshizuka, K., 2003. Selective recovery of lithiumfrom seawater using a novel MnO2 type adsorbent II—enhancement of lithium ionselectivity of the adsorbent. Ars Separatoria Acta 2, 97–106.

Kondás, J., Jandová, J., 2006. Lithium extraction from zinnwaldite wastes after gravitydressing of Sn-W ores. Acta Metall. Slovaca 12, 197–202.

Kondás, J., Jandová, J., Nemeckova, M., 2006. Processing of spent Li/MnO2 batteries toobtain Li2CO3. Hydrometallurgy 84, 247–249.

Koyanaka, Y., Yasuda, Y., 1977. Concentration of lithium in seawater by ion exchangeresin. Suiyo Kaishi 18, 523–526.

Krishnaswamy, S., 1979. India's Mineral Resources, 2nd edition. Oxford & Publishing Co.,New Delhi, p. 349.

Kuo, I., 2011. Is One Million Electric Cars by 2015 Too Ambitious? VB News February 10,2011. http://venturebeat.com/2011/02/10/one-million-electric-cars-2015/ (and http://www1.eere.energy.gov/vehiclesandfuels/pdfs/1_million_electric_vehicles_rpt.pdf).

Lain, M., 2002. Recycling of lithium ion cells and batteries. J. Power Sources 97, 736–738.Lee, C.K., Rhee, K.I., 2002. Preparation of LiCoO2 from spent lithium-ion batteries. J. Power

Sources 109, 17–21.Lee, C.K., Rhee, K.I., 2003. Reductive leaching of cathodic active materials from lithium ion

battery wastes. Hydrometallurgy 68, 5–10.Legers, L., 2008. The Trouble with Lithium — 2: Under the Microscope. Meridian Int. Res.,

(May 29, 2008).Li, Y.J., Zeng, G.S., 2011. Hydrometallurgical process for recovery and synthesis of LiCoO2

from spent lithium-ion batteries. IEEE, (978-1-4577-0290-7/11).Li, J., Li, X., Zhang, Y., Hu, Q., Wang, Z., Zhou, Y., Fu, F., 2009a. Study of spent battery

material leaching process. Trans. Nonferrous Metals Soc. China 19, 751–755.Li, J., Zhao, R., He, X., Liu, H., 2009b. Preparation of LiCoO2 cathode materials from spent

lithium-ion batteries. Ionics 15, 111–113.Li, L., Jing, G., Renjie, C., Feng, W., Shi, C., Xiaoxiao, Z., 2010a. Environmental friendly

leaching reagent for cobalt and lithium recovery from spent lithium-ion batteries.Waste Manag. 30, 2615–2621.

Li, L., Ge, J., Wu, F., Chen, R., Chen, S., Wu, B., 2010b. Recovery of cobalt and lithium from spentlithium ion batteries using organic citric acid as leachant. J. Hazard. Mater. 176, 288–293.

Li, L., Chen, R., Sun, F., Wu, F., Liu, J., 2011. Preparation of LiCoO2 films from spent lithiumion batteries by a combined recycling process. Hydrometallurgy 108, 220–225.

Li, L., Jun, L., Yang, R., Xiao, X.Z., Ren, J.C., Feng, W., Khalil, A., 2012. Ascorbic-acid-assisted re-covery of cobalt and lithium from spent Li-ion batteries. J. Power Sources 218, 21–27.

Lien, R.H., 1985. Recovery of lithium from a Montmorillonite-type clay. Bureau of Mines,Report of Investigations RI 8967.

Lin J.R., Fan C., Chang I.L., Shiu J.Y.; 2003. Clean process of recovering metals from wastelithium ion batteries. 2003. US Patent 6,514,311 B1.

Liu, Y.J., Hu, Q.Y., Li, X.H., Wang, Z.X., Guo, H.J., 2006. Recycle and synthesis of LiCoO2 fromincisors bound of Li-ion batteries. Trans. Nonferrous Metals Soc. China 16, 956–959.

Löf, G.O.G., Lewis, W.K., 1942. Lithium chloride from lepidolite. J. Ind. Eng. Chem. 34 (2),209–216.

Luong, V.T., Kang, D.J., An, J.W., Kim, M.J., Tran, T., 2013. Factor affecting the extraction oflithium from lepidolite. Hydrometallurgy 134–135, 54–61.

Luong, V.T., Kang, D.J., An, J.W., Dao, D.A., Kim, M.J., Tran, T., 2014. Iron sulphate roastingfor extraction of lithium from lepidolite. Hydrometallurgy 141, 8–16.

Lupi, C., Pasquali, M., 2003. Electrolytic nickel recovery from lithium-ion batteries. Miner.Eng. 16, 537–542.

Lupi, C., Pasquali, M., Dell'Era, A., 2005. Nickel and cobalt recycling from lithium-ionbatteries by electrochemical processes. Waste Manag. 25, 215–220.

Mcketta, J.J., 1988. Lithium and lithium compounds. Encyclopedia of Chemical Processingand Design vol. 28. Marcel Dekker.

McLaughlin W., Adams T.S., 1998. Li reclamation process. US Patent 5888463 A.Medina, L.F., El-Naggar, M.M.A.A., 1984. An alternative method for the recovery of lithium

from spodumene. Metall. Trans. B 15, 725–726.Miyai, Y., Ooi, K., Katoh, S., 1988. Recovery of lithium from seawater using a new type of

ion sieve adsorbent based on MgMn2O4. Sep. Sci. Technol. 23, 179–191.









































http://digicoll.manoa.hawaii.edu/techreports/PDF/USBM-691.pdf

http://digicoll.manoa.hawaii.edu/techreports/PDF/USBM-691.pdf







http://www.che.ncsu.edu/ILEET/phevs/lithium-availability/An_Abundance_of_Lithium.pdf

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http://venturebeat.com/2011/02/10/one-million-electric-cars-2015/

http://www1.eere.energy.gov/vehiclesandfuels/pdfs/1_million_electric_vehicles_rpt.pdf

http://www1.eere.energy.gov/vehiclesandfuels/pdfs/1_million_electric_vehicles_rpt.pdf


















































Mohr, S.H., Mudd, G.M., Giurco, D., 2010. Lithium resources and production: a criticalglobal assessment. Report Prepared for CSIRO Minerals Down Under Flagship.Department of Civil Engineering, Monash University Institute for Sustainable Futures,University of Technology, Sydney, Australia.

Mohr, S.H., Mudd, G.M., Giurco, D., 2012. Lithium resources and production: criticalassessment and global projections. Minerals 2, 65–84.

Nan, J., Han, D., Zuo, X., 2005. Recovery of metal values from spent lithium-ion batterieswith deposition and solvent extraction. J. Power Sources 152, 278–284.

Nazri, G.-A., Pistoia, G., 2004. Lithium Batteries— Science and Technology. Kluwer AcademicPublishers, Boston.

Nelli J.R., Arthur T.E., 1970. Recovery of lithium from bitterns. U.S. Patent 3,537,813.Nishihama, S., Onishi, K., Yoshizuka, K., 2011. Selective recovery process of lithium from seawa-

ter using integrated ion exchange methods. Solv. Extr. Ion Exch. 29, 421–431.Norton, J.J., Schlegel, D.M., 1955. Lithium resources of North America: a brief sum-

mary of the geology, distribution, production and reserves. Geological SurveyBulletin 1027-G.

Onishi, K., Nakamura, T., Nishihama, S., Yoshizuka, K., 2010. Synergistic solvent impregnat-ed resin for adsorptive separation of lithium ion. Ind. Eng. Chem. Res. 49, 6554–6558.

Ono, K., 2009. Lithium Resources and Demand — Lithium Supply & Markets Conference2009 (LSM'09) Report — JOGMEC Current Topics vol. 09–21.

Ooi, K., Miyami, Y., Katoh, S., Maeda, H., Abe, M., 1989. Topotactic Li+ insertion to λ-MnO2

in the aqueous phase. Langmuir 5, 150–157.Ooi, K., Miyai, Y., Sakakihara, J., 1991. Mechanism of Li+ insertion in spinel-typemanganese

oxide. Redox and ion-exchange reactions. Langmuir 7, 1167–1171.Paukov, E.I., Kovalevskaya, Y.A., Kiseleva, I.A., Shuriga, T.N., 2010. A low-temperature heat

capacity study of natural lithiummicas— heat capacity of zinnwaldite. J. Therm. Anal.Calorim. 99, 709–712.

Paulino, J.F., Busnardo, N.G., Afonso, J.C., 2008. Recovery of valuable elements from spentLi-batteries. J. Hazard. Mater. 150, 843–849.

Peterson J.A., Gloss G.H., 1959. Lithium values recovery process. US Patent 2,893, 828.Pourbaix, M., 1966. Atlas of Electrochemical EquilibriaEnglish edition by James A. Franklin

Pergamon Press, Oxford, London.Pranolo, Y., Zhang, W., Cheng, C.Y., 2010. Recovery of metals from spent lithium-ion

battery leach solutions with a mixed solvent extractant system. Hydrometallurgy102, 37–42.

Ritcey, G.M., Ashbrook, A.W., 1984. Solvent Extraction Principles and Applications toProcess Metallurgy, Part-1. Elsevier Science Publishers B.V., Amsterdam, TheNetherlands, pp. 99–100.

Roonwal, G.S., Shahriar, K., Ranjbar, H., 2005. Mineral Resources and Development. DayaPublishing House, Delhi, p. 150.

Schaller, W.T., 1937. Lithium minerals: industrial minerals and rocks. Am. Inst. Min. Met.Eng. 427.

Schweitzer, G.K., Pesterfield, L.L., 2010. The Aqueous Chemistry of the Elements, 1st edition.Oxford Univ. Press, London.

Shin, S.M., Kim, N.H., Sohn, J.S., Yang, D.H., Kim, Y.H., 2005. Development of a metal recoveryprocess from Li-ion battery wastes. Hydrometallurgy 79, 172–181.

Shuva, M.A.H., Kurny, A.S.W., 2013. Hydrometallurgical recovery of value metals fromspent lithium ion batteries. Am. J. Mater. Eng. Technol. 1 (1), 8–12.

Siame, E., Pascoe, R.D., 2011. Extraction of lithium frommicaceous waste from China clayproduction. Miner. Eng. 24, 1595–1602.

Sitando, O., Crouse, P.L., 2012. Processing of a Zimbabwean petalite to obtained lithiumcarbonate. Int. J. Miner. Process. 102–103, 45–50.

Sojka, R., 1998. Innovative recycling technologies for rechargeable batteries. Proc. of theFourth International Battery Recycling Congress, Hamburg, Germany.

Steinberg, M., Dang, V.D., 1975. Sustainable energy—without hot air. Chem. Abstr., BWLreport 20535-R 86.

Steinberg, M., Dang, V.D., 1976. Geological survey professional paper 1005. , p. 79.Strelow, F.W.E., Weinert, C.H.S.W., Van Der Walt, T.N., 1974. Separation of lithium from

sodium, beryllium and other elements by cation-exchange chromatography in nitricacid–methanol. Anal. Chim. Acta. 71, 123–132.

Sun, L., Qiu, K., 2011. Vacuumpyrolysis and hydrometallurgical process for the recovery ofvaluable metals from spent lithium-ion batteries. J. Hazard. Mater. 194, 378–384.

Sun, L., Qiu, K., 2012. Organic oxalate as leachant and precipitant for the recoveryof valuable metals from spent lithium-ion batteries. Waste Manag. 32 (8),1575–1582.

Suzuki, T., Nakamura, T., Inoue, Y., Niinae, M., Shibata, J., 2012. A hydrometallurgicalprocess for the separation of aluminium, cobalt, copper and lithium in acidic sulfatemedia. Sep. Purif. Technol. 98, 396–401.

Swain, B., Jeong, J., Lee, J.C., Lee, G.H., 2006. Separation of cobalt and lithium from mixedsulphate solution using Na-Cyanex 272. Hydrometallurgy 84, 130–138.

Swain, B., Jeong, J., Lee, J.C., Lee, G.H., Sohn, J.S., 2007. Hydrometallurgical process forrecovery of cobalt fromwaste cathodic active material generated during manufactur-ing of lithium ion batteries. J. Power Sources 167, 536–544.

Swain, B., Jeong, J., Lee, J.C., Lee, G.H., 2008. Development of process flow sheet for recov-ery of high pure cobalt from sulfate leach liquor of LIB industry waste: a mathemat-ical model correlation to predict optimum operational conditions. Sep. Purif. Technol.63, 360–369.

Swain, B., Jeong, J., Yoo, K., Lee, J.C., 2010. Synergistic separation of Co(II)/Li(I) for therecycling of LIB industry wastes by supported liquid membrane using Cyanex 272and DR-8R. Hydrometallurgy 101, 20–27.

Tahil, W., 2010. Howmuch lithium does a Li ion EV battery really need?Meridian Int. Res.,(March 5, 2010).

Takeuchi, T., 1980. Extraction of lithium from sea water with metallic aluminium. J. Nucl.Sci. Technol. 17 (12), 922–928.

Tanii T., Tsuzuki S., Honmura S., Kamimura T., Sasaki K., Yabuki M., Nishida K., 2003.Method for crushing cell. US Patent No. 6,524,737 B1.

Tedjar, F., 2013. Challenges for recycling advanced Li‐ion batteries. Proc. InternationalBattery Association (IBA2013), Barcelona, March 11‐15, 2013 (http://congresses.icmab.es/iba2013/images/Oral_26FEB.pdf).

Tedjar F., Foudraz J.C., 2007. Method for the mixed recycling of lithium-based anodebatteries and cells. US Patent 7,820,317 B2.

Tran, K.T., Luong, T.V., An, J.W., Kang, D.J., Kim,M.J., Tran, T., 2013. Recovery ofmagnesium fromUyuni salar brine as high purity magnesium oxalate. Hydrometallurgy 138, 93–99.

U.S. Bureau of Mines and U.S. Geological Survey, 1980. Principles of a Resource/ReserveClassification for Minerals: U.S. Geological Survey Circular 831. http://pubs.usgs.gov/circ/1980/0831/report.pdf.

U.S. Bureau of Mines and U.S. Geological Survey, 1986. Principles of a Resource/ReserveClassification for Minerals. BuMInes Miner. Commod. Sum. 1986, pp. 182–185.

U.S. Geological Survey, 2005. Mineral Commodity Summaries, (January).U.S. Geological Survey, 2009. Mineral Commodity Summaries, (January).U.S. Geological Survey, 2010. Mineral Commodity Summaries. U.S. Geological

Survey, Reston, VA, USA.U.S. Geological Survey, 2012. Mineral Commodity Summaries. January 2012. http://

minerals.usgs.gov/minerals/pubs/commodity/lithium/mcs-2012-lithi.pdf.U.S. Geological Survey, 2013. Mineral Commodity Summaries. January 2013. http://

minerals.usgs.gov/minerals/pubs/commodity/lithium/mcs-2013-lithi.pdf.Um, N., Hirato, T., 2012. A study on lithium recovery from seawater: separation of lithium

from hydrochloric acid solutions containing CaCl2, MgCl2, MnCl2, NaCl, KCl and LiCl.Zero-Carbon Energy Kyoto (2012). In: Yao, T. (Ed.), Green Energy and Technology,pp. 149–152.

Umeno, A., Miyai, Y., Takagi, N., Chitrakar, R., Sakane, K., Ooi, K., 2002. Preparation andadsorptive properties of membrane-type adsorbents for lithium recovery fromseawater. Ind. Eng. Chem. Res. 41 (17), 4281–4287.

Vu, H., Bernardi, J., Jandová, J., Vaculíková, L., Goliáš, V., 2013. Lithium and rubidiumextraction from zinnwaldite by alkali digestion process: sintering mechanism andleaching kinetics. Int. J. Miner. Process. 123, 9–17.

Vyas, M.H., Sanghavi, J.R., Seshadri, K., 1975. Lithium extraction from Indian lepidoliteores. Res. Ind. 20, 68–70.

Wajima, T., Munakata, K., Uda, T., 2012. Adsorption behavior of lithium fromseawater using manganese oxide adsorbent. Plasma Fusion Res. 7. The JapanSociety of Plasma Science and Nuclear Fusion Research, pp. 2405021-1–2405021-4.

Wakihara, M., Yamamoto, O., 1998. Lithium Ion Batteries Fundamentals and Performance.Kodansha Ltd., Tokyo 1998.

Wang, R.C., Lin, Y.C., Wu, S.H., 2009. A novel recovery process of metal values from thecathode active materials of the lithium-ion secondary batteries. Hydrometallurgy99, 194–201.

Wang, J., Chen, M., Chen, H., Luo, T., Xu, Z., 2012. Leaching study of spent Li-ionbatteries. Proc. Environ. Sci. (The 7th Int. Conf. waste manage. Technol.) 16,pp. 443–450.

Yaksic, A., Tilton, J.E., 2009. Using the cumulative availability curve to assess the threat ofmineral depletion: the case of lithium. Resour. Policy 34, 185–194.

Yamaji, Y., Dodbiba, G., Matsuo, S., Okaya, K., Shibayama, A., Fujita, T., 2011. A novel flowsheet for processing of used lithium-ion batteries for recycling. Resour. Process. 58(1), 9–13.

Yan, Q., Li, X., Wang, Z., Wu, X., Wang, J., Guo, H., Hu, Q., Peng, W., 2012a. Extraction oflithium from lepidolite by sulfation roasting and water leaching. Int. J. Miner. Process.110–111, 1–5.

Yan, Q., Li, X., Wang, Z., Wu, X., Guo, H., Hu, Q., Peng, W., Wang, J., 2012b. Extraction ofvaluable metals from lepidolite. Hydrometallurgy 116–118.

Yan, Q., Li, X., Yin, Z., Wang, Z., Guo, H., Peng, W., Hu, Q., 2012c. A novel process forextracting lithium from lepidolite. Hydrometallurgy 121–124, 54–59.

Yan, Q., Li, X., Wang, Z., Wang, J.X., Guo, H., Hu, Q., Peng, W., Wu, X.F., 2012d. Extraction oflithium from lepidolite using chlorination roasting–water leaching process. Trans.Nonferrous Metals Soc. China 22 (7), 1753–1759.

Yoshinaga, T., Kawano, K., Imoto, H., 1986. Basic study on lithium recovery from lithiumcontaining solution. Bull. Chem. Soc. Jpn. 59, 1207–1213.

Yoshizuka, K., Fukui, K., Inoue, K., 2002. Selective recovery of lithium from seawater usinga novel MnO2 type adsorbent. Ars Separatoria Acta 1, 79–86.

Yoshizuka, K., Kitajou, A., Holba, M., 2006. Selective recovery of lithium from seawater using anovel MnO2 type adsorbent III—benchmark evaluation. Ars Separatoria Acta 4, 78–85.

Zelikman, A.N., Krein, O.E., Samsonov, G.V., 1996. Metallurgy of Rare MetalsTranslatedfrom Russian & published by NASA and National Science Foundation, WashingtonD.C., p. 458.

Zhang, P., Yokoyama, T., Itabashi, O., Suzuki, T.M., Inoue, K., 1998. Hydrometallurgicalprocess for recovery of metal values from spent lithium-ion secondary batteries.Hydrometallurgy 47, 259–271.

Zhao, J.M., Shen, X.Y., Deng, F.L., Wang, F.C., Wu, Y., Liu, H.Z., 2011. Synergisticextraction and separation of valuable metals from waste cathodic material oflithium ion batteries using Cyanex272 and PC-88A. Sep. Purif. Technol. 78(3), 345–351.

Zhao, Z., Si, X., Liu, X., He, L., Liang, X., 2013. Li extraction from high Mg/Li ratio brine withLiFePO4/FePO4 as electrode materials. Hydrometallurgy 133, 75–83.

Zhu, S., He,W., Li, G., Zhou, X., Huang, J., Zhang, X., 2011. Recovering copper from spent lithiumionbatterybyamechanical separationprocess. Proc. InternationalConferenceonMaterialsfor Renewable Energy & Environment (ICMREE), Shanghai vol. 2, pp. 1008–1012.

Zhu, S.G., He, W.Z., Li, G.M., Zhou, X., Zhang, X.J., Huang, J.W., 2012. Recovery of Coand Li from spent lithium-ion batteries by combination method of acidleaching and chemical precipitation. Trans. Nonferrous Metals Soc. China 22,2274–2281.





















































































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